Lecture Notes On Chemical Methods of Extractive Metallurgy

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Notes on Chemical Methods of mineral Extraction

EXTRACTIVE METALLURGY II
INTRODUCTION
Extractive metallurgy is the process of extracting valuable metals from their natural sources (ore) and refining the extracted raw metals into a purer form for practical use. The field is an applied science, covering all aspects of the physical and chemical processes used to produce mineral-containing and metallic materials, sometimes for direct use as a finished product, but more often in a form that requires further physical processing which is generally the subject of physical metallurgy, ceramics and other disciplines within the broad field of materials science.
The field of extractive metallurgy encompasses many specialty sub-disciplines, each concerned with various physical and chemical processes that are steps in an overall process of producing a particular material. These specialties are generically grouped into the categories of mineral processing, hydrometallurgy, pyrometallurgy, and electrometallurgy. The distinction among these groups of sub-disciplines within extractive metallurgy is far from clear, and many commercially important metallurgical processes involve considerable overlap.
The theoretical basis of extractive metallurgy is underpinned by the more general sciences of physics, chemistry, and geology. Additionally, the practice of extractive metallurgy nearly always involves contributions from other scientific fields such as analytical chemistry and mineralogy.
Extractive metallurgists are interested in three primary streams: feed, concentrate (valuable metal oxide/sulfide), and tailings (waste). After mining, large pieces of the ore feed are broken through crushing and/or grinding in order to obtain particles small enough where each particle is either mostly valuable or mostly waste. Concentrating the particles of value in a form supporting separation enables the desired metal to be removed from waste products.
Mining may not be necessary if the ore body and physical environment are conducive to leaching. Leaching dissolves minerals in an ore body and results in an enriched solution. The solution is collected and processed to extract valuable metals.
Ore bodies often contain more than one valuable metal. Tailings of a previous process may be used as a feed in another process to extract a secondary product from the original ore. Additionally, a concentrate may contain more than one valuable metal. That concentrate would then be processed to separate the valuable metals into individual constituents.
PRINCIPLES OF EXTRACTION
The extraction of pure metal from its ore involves several physical and chemical methods. The suitability of the method depends upon the nature of ore, the properties of metals and local conditions. The extractive metallurgy of a metal involves the following operations:
1. Crushing of ores
2. Concentration of ores
3. Reduction
4. Purification of ores
ORE REDUCTION METHODS
Some of the methods commonly used to get free metal from ore are given below:
PYROMETALLURGICAL ORE REDUCTION METHODS
1. Smelting
Smelting is the most common pyrometallurgical process and involves the application of heat and a chemical reducing agent to decompose ore, driving off other elements as gasses or slag and leaving just the metal behind. The reducing agent is commonly a source of carbon such as coke. The carbon removes oxygen from the ore as carbon monoxide, leaving behind elemental metal. Other reducing agents in use today include H2, CO, water, Na, K, Mg and Al.
Iron Smelting:
Zone 1 (<950°C), reduction of Fe2O3, Fe3O4 takes place:
3Fe2O3(s) + CO(g) → 2Fe3O4(s) + CO2(g)
Fe3O4(s) + CO(g) → 3FeO(s) + CO2(g)
Zone 2 (950-1,000°C), chemically reserve zone, FeO is in equilibrium with the gaseous phase:
FeO(s) + CO(g) ↔ Fe(l) + CO2(g)
Zone 3 (950-1,050°C), the reduction of FeO by rising CO gas takes place:
FeO(s) + CO(g) → Fe(l) + CO2(g)
Zone 4 (>1,000-1,050°C), direct reduction of FeO by C takes place:
FeO(s) + C(s) → Fe(l) + CO(g)
Raceway zone (below Zone 4) (1,050°C), conversion of CO2 to CO:
C(s) + O2(g) → CO2(g)
CO2(g) + C(s) → 2CO(g) (Boudward reaction)
2C(s) + O2(g) ↔ 2CO(g) (Overall reaction)
The gangue consists of SiO2, Al2O3, P, S bearing minerals. Their removal proceeds thus:
CaO(s) + SiO2(s) → CaSiO3(l) (1,200°C)------slag
CaO(s) + P2O5(s) → Ca3(PO4)2(l) (1,200°C)--slag
Fig. 1: Reaction zones in a blast furnace
Flash smelting
Flash smelting is a smelting process for sulfur-containing ores including chalcopyrite. The process was developed in 1949 in Finland for smelting copper ore. It has also been adapted for nickel and lead production. The process uses the autogenic principle by using the energy contained in the sulfur and iron for melting the ore. In the process dried and powdered ore is discharged from a nozzle into a fluidized bed reactor fed with oxygen. The reduced metal melts, and drops to the bottom of a settling chamber. The flotation produces a large effective surface area of fine-grained concentrate particles. The process makes smelting more energy efficient and environmentally friendly. Sulfur is released mainly in its solid form, thus reducing atmospheric pollution. The process is today used for 50% of the world’s primary copper production. The other 50% is mainly produced from oxide ores, where the process cannot be applied.
Copper smelting
Ores of Cu
chalcopyrite (CuFeS2), chalcocite (Cu2S), covellite (CuS), bornite (2Cu2S·CuS·FeS), tetrahedrite (Cu3SbS3+x(Fe,Zn)6Sb2S9), malachite (CuCO3·Cu(OH)2), azurite (2CuCO3·Cu(OH)2), cuprite (Cu2O), chrysocolia (CuO·SiO2·2H2O) and tennantite (Cu12As4S13).
Roasting
The copper ore concentrate is placed in a roaster and is partially oxidized to produce “calcine” and sulphur dioxide gas:
2CuFeS2(s) + 3O2(g) → 2FeO(s) + 2CuS(s) + 2SO2(g)
Roasting is no longer in common use as direct smelting is now favoured.
Smelting
The calcine is then mixed with silica and coke and smelted in an exothermic reaction at 1200°C (above the melting point of copper, but below that of iron and silica) to form a liquid called “copper matte”. The high temperature allows reactions to proceed rapidly, and allow the matte and slag to melt, so they can be tapped out of the furnace. In copper recycling, this is the point where scrap copper is introduced. Several reactions occur:
FeO(s) + SiO2(s) → FeSiO3(l)
In a parallel reaction, iron sulfide is converted to slag:
2FeS(s) + 3O2 + SiO2(s) → 2FeSiO3(l) + 2SO2(g)
The slag is discarded or reprocessed to recover any remaining copper. The sulphuric acid is captured for use in earlier leaching processes.
Conversion to blister
The matte, which is produced in the smelter, contains around 70% copper primarily as copper sulfide as well as iron sulfide. The sulphur is removed at high temperature as sulphur dioxide by blowing air through molten matte:
2CuS + 3O2 → 2CuO + 2SO2
CuS + O2 → Cu + SO2
In a parallel reaction the iron sulfide is converted to slag:
2FeS + 3O2 → 2FeO + 2SO2
2FeO + 2SiO2 → 2FeSiO3
The purity of this product is 98%, it is known as blister because of the broken surface created by the escape of sulfur dioxide gas as the copper pigs or ingots are cast. By-products generated in the process are sulfur dioxide and slag. The sulfur dioxide is captured for use in earlier leaching processes.
Reduction
The blistered copper is put into an anode furnace (a furnace that uses the blister copper as anode) to get rid of most of the remaining oxygen. This is done by blowing natural gas through the molten copper oxide. When this flame burns green, indicating the copper oxidation spectrum, the oxygen has mostly been burned off. This creates copper at about 99% pure. The anodes produced from this are fed to the electrorefinery.
2. Self-reduction process
The sulphide ores of less electropositive metals like Hg, Pb, Cu, etc. are heated in air to convert part of sulphide ore into oxide, which then reacts with the remaining sulphide to give the metal and sulphur dioxide.
2HgS + 3O2 → 2HgO + 2SO2
2HgO + HgS → 3Hg + SO2
3. Roasting
Roasting is a step of the processing of certain ores. It is a pyrometallurgical process involving gas–solid reactions at elevated temperatures with the goal of purifying the metal component(s). Often before roasting, the ore has already been partially purified, e.g. by froth floatation. The concentrate is mixed with other materials to facilitate the process. The technology is useful but is also a serious source of air pollution.
Roasting consists of thermal gas–solid reactions, which can include oxidation, reduction, chlorination, sulfation, and pyrohydrolysis. In roasting, the ore or ore concentrate is treated with very hot air. This process is generally applied to sulphide minerals. During roasting, the sulfide is converted to an oxide, and sulfur is released as sulfur dioxide, a gas. For the ores Cu2S (chalcocite) and ZnS (sphalerite), balanced equations for the roasting are:
2Cu2S + 3O2 → 2Cu2O + 2SO2
2ZnS + 3O2 → 2ZnO + 2SO2
The gaseous product of sulfide roasting, sulfur dioxide (SO2) is often used to produce sulfuric acid. Many sulfide minerals contain other components such as arsenic that are released into the environment.
Up until the early 20th century, roasting was started by burning wood on top of ore. This would raise the temperature of the ore to the point where its sulfur content would become its source
of fuel, and the roasting process could continue without external fuel sources. Early sulfide roasting, was practiced in this manner, in "open hearth" roasters, which were manually stirred (a practice referred to as "rabbling") using rake-like tools to expose unroasted ore to oxygen as the reaction proceeded.
4. Calcination
Calcination (also referred to as calcining) is a thermal treatment process in the absence of air applied to ores and other solid materials to bring about a thermal decomposition, phase transition, or removal of a volatile fraction. The calcination process normally takes place at temperatures below the melting point of the product materials. Calcination is not the same process as roasting. In roasting, more complex gas–solid reactions take place between the furnace atmosphere and the solids.
The process of calcination derives its name from the Latin calcinare (to burn lime) due to its most common application, the decomposition of calcium carbonate (limestone) to calcium oxide (lime) and carbon dioxide (Fig. 2), in order to produce cement. The product of calcination is usually referred to in general as "calcine," regardless of the actual minerals undergoing thermal treatment. Calcination is carried out in furnaces or reactors (sometimes referred to as kilns or calciners) of various designs including shaft furnaces, rotary kilns, multiple hearth furnaces, and fluidized bed reactors.
Fig. 2: An oven for calcination of limestone
Examples of calcination processes include the following:
 decomposition of carbonate minerals, as in the calcination of limestone to drive off carbon dioxide;
 decomposition of hydrated minerals, as in the calcination of bauxite and gypsum, to remove crystalline water as water vapor;
 decomposition of volatile matter contained in raw petroleum coke;
 heat treatment to effect phase transformations, as in conversion of anatase to rutile or devitrification of glass materials
 removal of ammonium ions in the synthesis of zeolites
Calcination reactions usually take place at or above the thermal decomposition temperature (for decomposition and volatilization reactions) or the transition temperature (for phase transitions). This temperature is usually defined as the temperature at which the standard Gibbs free energy for a particular calcination reaction is equal to zero. For example, in limestone calcination, a decomposition process, the chemical reaction is
CaCO3 → CaO + CO2(g)
The standard Gibbs free energy of reaction is approximated as ΔG°r = 177,100 − 158 T (J/mol). The standard free energy of reaction is zero in this case when the temperature, T, is equal to 1121 K, or 848 °C.
In some cases, calcination of a metal may results in oxidation of the metal, leading to it gaining weight, e.g. Pb and Sn.
5. Liquation
Liquation is a metallurgical method for separating metals from an ore or alloy. The material must be heated until one of the metals starts to melt and drain away from the other and can be collected. This method was largely used to remove lead containing silver from copper, but it can also be used to remove antimony minerals from ore, and to refine tin.
Separating copper and silver
The first known use of Liquation on a large scale was in Germany in the mid-15th century. Liquation requires that the silver-rich copper first be melted with approximately three times its weight in lead, as silver has a greater affinity with lead most of the silver would end up within this rather than the copper. If the copper is assayed and found to contain too little silver for
liquation to be financially viable (around 0.31% is the minimum required) it is melted and allowed to settle so that much of the silver sinks towards the bottom. The ‘tops’ are then drawn off and used to produce copper while the silver-rich ‘bottoms’ are used in the liquation process. The copper-lead alloy created can be tapped off and cast into large plano-convex ingots known as ‘liquation cakes’. As the metals cool and solidify the copper and the silver-containing lead separate as they are immiscible with each other.
The ratio of lead to copper in these cakes is an important factor for the process to work efficiently. Three parts copper to 8-12 parts lead is recommended. The copper must be assayed to accurately determine how much silver it contains, for copper rich in silver the top end of this ratio was used to make sure the maximum amount of silver possible would end up within the lead. However there also needs to be enough copper to allow the cakes to keep their shape once most of the lead has drained away, too much copper and it would trap some of the lead within and the process would be very inefficient.
The size of these cakes remained consistent from 1556 to the 19th century when the process became obsolete. They were usually between 2½ to 3½ inches (6.4 to 8.9 cm) thick, about 2 feet (0.61m) in diameter and weighed from 225 to 375 lbs (102 kg to 170 kg). This consistency is not without reason as the size of the cakes is very important to the smooth running of the liquation process. If the cakes are too small, the product would not be worth the time and costs spent on the process, if they are too large then the copper would begin to melt before the maximum amount of lead has drained away.
The cakes are heated in a liquation furnace, usually four or five at once, to a temperature above the melting point of lead (327°C), but below that of copper (1084°C), so that the silver-rich lead melts and flows away. As the melting point of lead is so low a high temperature furnace is not required and it can be fuelled with wood. It is important that this takes place in a reducing atmosphere, i.e. one with little oxygen, to avoid the lead oxidising, the cakes are therefore well covered by charcoal and little air is allowed into the furnace. It is impossible to stop some of the lead oxidising however and this drops down and forms spiky projections known as ‘liquation thorns’ in the channel underneath the hearth.
The older and relatively simple method of cupellation can then be used to separate the silver from the lead. If the lead is assayed and found not to contain enough silver to make the cupellation process worthwhile it is reused in liquation cakes until it has sufficient silver.
The ‘exhausted liquation cakes’ which still contain some lead and silver are ‘dried’ in a special furnace which is heated to a higher temperature under oxidising conditions. This is essentially just another stage of liquation and most of the remaining lead is expelled and oxidised to form
liquation thorns, though some remains as lead metal. The copper can then be refined to remove other impurities and produce copper metal.
Waste products can be reused to produce new liquation cakes to try and minimise loss of metals, especially silver. The waste products are mostly in the form of liquation thorns from the liquation and the drying process but there are also some slags produced.
HYDROMETALLURGICAL ORE REDUCTION METHODS
1. Leaching
Leaching is a widely used extractive metallurgy technique which converts metals into soluble salts in aqueous media. Compared to pyrometallurgical operations, leaching is easier to perform and much less harmful, because no gaseous pollution occurs. Drawbacks of leaching are the highly acidic and in some cases toxic residual effluent, and its lower efficiency caused by the low temperatures of the operation, which dramatically affect chemical reaction rates.
There are a variety of leaching processes, usually classified by the types of reagents used in the operation. The reagents required depend on the ores or pretreated material to be processed. A typical feed for leaching is either oxide or sulfide.
For material in oxide form, a simple acid leaching reaction can be illustrated by the zinc oxide leaching reaction:
ZnO + H2SO4 → ZnSO4 + H2O
In this reaction solid ZnO dissolves, forming soluble zinc sulfate.
In many cases other reagents are used to leach oxides. For example, in the metallurgy of aluminium, aluminium oxide is subject to leaching by alkali solutions:
Al2O3 + 3H2O + 2NaOH → 2NaAl(OH)4
Leaching of sulfides is a more complex process due to the refractory nature of sulfide ores. It often involves the use of pressurized vessels, called autoclaves. A good example of the autoclave leach process can be found in the metallurgy of zinc. It is best described by the following chemical reaction:
2ZnS + O2 + 2H2SO4 → 2ZnSO4 + 2H2O + 2S
This reaction proceeds at temperatures above the boiling point of water, thus creating a vapor pressure inside the vessel. Oxygen is injected under pressure, making the total pressure in the autoclave more than 0.6 MPa.
The leaching of precious metals such as gold can be carried out with cyanide or ozone under mild conditions. A lixiviant is a liquid medium used in hydrometallurgy to selectively extract the desired metal from the ore or mineral. It assists in rapid and complete leaching. The metal can be recovered from it in a concentrated form after leaching. Lixiviant in a solution may be acidic or basic in nature. The most commonly used acid is H2SO4.
Heap leaching
Heap leaching is an industrial mining process to extract precious metals, copper, uranium, and other compounds from ore via a series of chemical reactions that absorbs specific minerals and then re-separate them after their division from other earth materials. Comparable to in situ mining, heap leach mining differs in that it uses a liner to place amounts of ore on, then adds the chemicals via drip systems to the ore, whereas in situ leaching lacks these pads and pulls pregnant solution up to obtain the minerals. This method is only slightly friendlier environmentally, however, and has still seen copious amounts of negative feedback from both environmentalists and health experts in the past twenty or more years. Since its original peak of popularity in the 1970s, the heap leach technique has been applied throughout the earth, but due to recent increases in negative environmental impact assessments has received more discussion regarding rehabilitation than perpetuation of these types of techniques. Yet this method continues to be a profit-earning endeavor for many mining companies across the globe.
The process has ancient origins; one of the classical methods for the manufacture of copperas (iron sulfate) was to heap up iron pyrite and collect the leachate from the heap, which was then boiled with iron to produce iron sulfate
The mined ore is usually crushed into small chunks and heaped on an impermeable plastic and/or clay lined leach pad where it can be irrigated with a leach solution to dissolve the valuable metals. While sprinklers are occasionally used for irrigation, more often operations use drip irrigation to minimize evaporation, provide more uniform distribution of the leach solution, and avoid damaging the exposed mineral. The solution then percolates through the heap and leaches both the target and other minerals. This process, called the "leach cycle," generally takes from one or two months for simple oxide ores (e.g., most gold ores) to two years (for nickel laterite ores). The leach solution containing the dissolved minerals is then collected, treated in a process plant to recover the target mineral and in some cases precipitate other minerals, and then recycled to the heap after reagent levels are adjusted. Ultimate recovery of the target mineral can range from 30% of contained (run-of-mine dump leaching sulfide copper ores) to over 90% for the easiest to leach ores (some oxide gold ores).
In recent years, the addition of an agglomeration drum has improved on the heap leaching process by allowing for a more efficient leach. The rotary drum agglomerator works by taking the crushed ore fines and agglomerating them into more uniform particles. This makes it much easier for the leaching solution to percolate through the pile, making its way through the channels between particles.
The addition of an agglomeration drum also has the added benefit of being able to pre-mix the leaching solution with the ore fines, to achieve a more concentrated, homogeneous mixture, and allowing the leach to begin prior to the heap.
Precious metals
The crushed ore is irrigated with a dilute alkaline cyanide solution. The solution containing the dissolved precious metals ("pregnant solution") continues percolating through the crushed ore until it reaches the liner at the bottom of the heap where it drains into a storage (pregnant solution) pond. After separating the precious metals from the pregnant solution, the dilute cyanide solution (now called "barren solution") is normally re-used in the heap-leach-process or occasionally sent to an industrial water treatment facility where the residual cyanide is treated and residual metals are removed. In very high rainfall areas, such as the tropics, in some cases there is surplus water that is then discharged to the environment, after treatment, posing possible water pollution if treatment is not properly carried out.
The production of one gold ring through this method can generate 20 tons of waste material.
During the extraction phase, the gold ions form complex ions with the cyanide:
Au+(s) + 2CNˉ(aq) → Au(CN)2ˉ(aq)
Recuperation of the gold is readily achieved with a redox-reaction:
2Au(CN)2ˉ(aq) + Zn(s) → Zn(CN)4ˉ(aq) + 2Au(s)
The most common methods to remove the gold from solution are either using activated carbon to selectively absorb it, or the Merrill-Crowe process where zinc powder is added to cause a precipitation of gold and zinc. The fine product can be either doré (gold-silver bars) or zinc-gold sludge that is then refined elsewhere.
Copper Ores
The method is similar to the cyanide method, above, except sulfuric acid is used to dissolve copper from its ores. The acid is recycled from the solvent extraction circuit and reused on the leach pad. A byproduct is iron (II) sulfate, jarosite, which is produced as a byproduct of leaching pyrite, and sometimes even the same sulfuric acid that is needed for the process. Both oxide
and sulfide ores can be leached, though the leach cycles are much different and sulfide leaching requires a bacterial or "bio-leach" component. The largest copper heap leach operations are in Chile, Peru, and the southwestern United States.
Although the heap leaching is a low cost-process, it normally has recovery rates of 60-70%, although there are exceptions. It is normally most profitable with low-grade ores. Higher-grade ores are usually put through more complex milling processes where higher recoveries justify the extra cost. The process chosen depends on the properties of the ore.
The final product is cathode copper.
Nickel Ores
The method is an acid heap leaching method like that of the copper method in that it utilises sulfuric acid instead of cyanide solution to dissolve the target minerals from crushed ore. The amount of sulfuric acid required is much higher than for copper ores (as high as 1,000 kg of acid per tonne of ore, but 500 kg is more common.) Nickel recovery from the leach solutions is much more complex than for copper and requires various stages of iron and magnesium removal, and the process produces both leached ore residue ("ripios") and chemical precipitates from the recovery plant (principally iron oxide residues, magnesium sulfate and calcium sulfate) in roughly equal proportions. Thus, a unique feature of nickel heap leaching is the need for a tailings disposal area.
The final product can be nickel hydroxide precipitates (NHP) or mixed metal hydroxide precipitates (MHP), which are then subject to conventional smelting to produce metallic nickel.
The method was originally patented by Australian miner BHP Billiton and is being commercialized by Cerro Matoso S.A. in Colombia (a wholly owned subsidiary of BHP Billiton), Vale in Brazil, and European Nickel PLC for the rock laterite deposits of Turkey, Talvivaara mine in Finland, Balkans, and the Philippines. There currently are no operating commercial scale nickel laterite heap leach operations, but there is a sulfide HL operating in Finland.
Uranium Ores
Heap leaching of uranium ores is similar to copper oxide heap leaching, also using dilute sulfuric acid. Rio Tinto is commercializing this technology in Namibia and Australia, the French nuclear power company Areva in Niger (two mines) and Namibia, and several other companies are studying its feasibility.
The final product is yellowcake and requires significant further processing to produce fuel-grade feed.
Heap Leaching Apparatus
While most mining companies have shifted from a previously accepted sprinkler method to the percolation of slowly dripping choice chemicals (cyanide or sulfuric acid) closer to the actual ore bed, heap leach pads have not changed too much throughout the years. There are still four main categories of pads: conventional, dump leach, Valley Fills, and on/off pads. Typically, each pad only has a single, geomembrane liner for each pad, with a minimum thickness of 1.5mm (usually it is thicker).
The simplest in design, conventional pads are used for mostly flat or gentle areas and hold thinner layers of crushed ore. Dump leach pads hold more ore and can usually handle a less flat terrain. Valley Fills are pads situated at valley bottoms or levels that can hold everything falling into it. On/off pads involve the use of putting significantly larger loads on the pads, and removing it and reloading it after every cycle.
Many of these mines, which previously had digging depths of about 15 meters, are digging deeper than ever before to mine materials (approximately 50 meters, sometimes more), which means that, in order to accommodate all of the ground being displaced, pads will have to hold higher weights from more crushed ore being contained in a smaller area. With that increase in buildup comes in potential for decrease in yield or ore quality, as well as potential either weak spots in the lining or areas of increased pressure buildup. This build up still has the potential to lead to punctures in the liner. As of 2004 cushion fabrics, which could reduce potential punctures and their leaking, were still being debated due to their tendency to increase risks if too much weight on too large a surface was placed on the cushioning. In addition, some liners, depending on their composition, may react with salts in the soil as well as acid from the chemical leaching to affect the successfulness of the liner. This can be amplified over time.
Heap Leaching Environmental Concerns
Heap leach mining works well for high concentrations of less ores, as less Earth needs to be ground onto leach pads in order to extract the same amount of materials. While yield is usually approximately 60-70%, there are significant amounts of damage to the surface environment. Yet the amount of overall harm caused by heap leaching is often lower than more traditional techniques, reducing costs to the process. It also requires less energy consumption to use these methods, which many consider to be an environmental alternative.
In some cases, waste materials from this process are transported to a facility to be treated. However, after a month (or more) of setting in a mat, there is often still more wait time involved in recovering the chemicals—both pregnant and excess—thus allowing for additional chemicals to potentially leach out of the pad into the soil below. This could possibly cause damage to the environment, which has a chance at contaminating surrounding bodies of water.
All in all, it is only slightly more environmentally friendly than in-situ mining, which involves leaving chemicals directly into the soil before pulling pregnant chemicals up. Still, most soil in heap leaching is seldom replaced or additionally treated. Evidence has also been found to show that there increased levels of erosion in mining sites with open chemicals, including these heap leach mines, which could be exacerbated through natural phenomena like storms and wind, or more serious occurrences like earthquakes.
The threat to ecosystem composition and biodiversity has been commented on repeatedly in terms of these mines, and it is noted that, while they do have a higher yield, they also have tendencies to accumulate wear and tear from being outside, and pose a threat to the immediate environment by crushing and dumping dirt that would otherwise have been left untouched. As noted earlier, with the reduction of readily available rare earth minerals, there has been increased in the amount of ore piled onto these pads, suggesting that there may come a time when the amount of ore dumped is not worth the amount of returned mineral collected. Therefore, alternatives need to be considered in the near future. Currently, depths are being mined faster than research can provide information about the effects of more ore on the system.
There is also very little study for long-term viability of liners, as this type of mining and the increased ore depths are still a relatively new field, especially given the changes in depths that have been put into practice. With the increase in weight, pressure, and chemicals put on this method of mining, as well as the already small level of knowledge regarding long term benefits, it is difficult to predict the extent of damage from previous leaks, as well as the durability of present day pads and mining sites.
Heap Leaching Legal Claims
Mining laws have been lobbied against in the United States for the past few decades, and many are only beginning to hear the environmentalist argument. The US General Mining Law of 1872 previous gave liberal rights to miners in terms of establishing and exploring claims, yet did not require any sort of environmental rehabilitation aspect to its process. Today, this is being disputed due to the number of environmental problems being found in heap leach sites, as well as the increase in scientific knowledge that could make mining more efficient and less costly as well. There has been much debate going into the levels of revision of the U.S. General Mining Law of 1872, including whether rehabilitation measures should be added to a decrease in the amount of liberal rights given to mining companies. Australia has already addressed much of this with the increased amounts of impact and externality knowledge required in any mining proposal endeavor.
However, as seen with many case studies, a simple way around these measures is the privatization of the land to be mined. In this case, environmental standards need to be made priority, as their effects spread beyond simple legal boundaries and into the ecosystems present at each mine, many of which are affecting poorer people less likely to speak up for their health and their environment.
Heap Leaching Cultural and Social Concerns
With the rise of the environmentalist movement has also come an increased appreciation for social justice, and mining has showed similar trends lately. Societies located near potential mining sites are at increased risk to be subjected to injustices as their environment is affected by the changes made to mined lands—either public or private—that could eventually lead to problems in social structure, identity, and physical health. Many have argued that by cycling mine power through local citizens, this disagreement can be alleviated, since both interest groups would have shared and equal voice and understanding in future goals. However, it is often difficult to match corporate mining interests with local social interests, and money is often a deciding factor in the successes of any disagreements. If communities are able to feel like they have a valid understanding and power in issues concerning their local environment and society, they are more likely to tolerate, and indeed encourage, the positive benefits that come with mining, as well as more effectively promote alternative methods to heap leach mining using their intimate knowledge of the local geography. Through increased dialogue and environmental legislation, many corporations and citizens hope to bridge the gap between interests in order to obtain the rare natural resources that most people depend on in daily life.
Dump leaching
Dump leaching is an industrial process to extract precious metals and copper from ores.
Dump leaching is similar to heap leaching, however in the case of dump leaching ore is taken directly from the mine and stacked on the leach pad without crushing where, in the case of gold and silver, the dump is irrigated with a dilute cyanide solution that percolates through the ore to dissolve gold and silver. The solution containing gold and silver exits the base of the dump, is collected and precious metals extracted. The resultant barren solution is recharged with additional cyanide and returned to the dump.
This method of leaching is usually suitable for low grade ores because it is very low cost. However, it operates with slow kinetics and may take up about 1 to 2 years to extract 50% of the desired mineral.
Tank and Vat leaching
In metallurgical processes tank leaching is a hydrometallurgical method of extracting valuable material (usually metals) from ore.
Tank vs. vat leaching
Tank leaching is usually differentiated from vat leaching on the following factors:
1. In tank leaching the material is ground sufficiently fine to form a slurry or pulp, which can flow under gravity or when pumped. In vat leaching typically a coarser material is placed in the vat for leaching, this reduces the cost of size reduction;
2. Tanks are typically equipped with agitators, baffles, gas introduction equipment designed to maintain the solids in suspension in the slurry, and achieve leaching. Vats usually do not contain “internal” equipment;
3. Tank leaching is typically continuous, while vat leaching is operated in a batch fashion, this is not always the case, and commercial processes using continuous vat leaching have been tested;
4. Typically the retention time required for vat leaching is more than that for tank leaching to achieve the same percentage of recovery of the valuable material being leached;
In a tank leach the slurry is moved, while in a vat leach the solids remain in the vat, and solution is moved.
Tank and vat leaching involves placing ore, usually after size reduction and classification, into large tanks or vats at ambient operating conditions containing a leaching solution and allowing the valuable material to leach from the ore into solution.
In tank leaching the ground, classified solids are already mixed with water to form a slurry or pulp, and this is pumped into the tanks. Leaching reagents are added to the tanks to achieve the leaching reaction. In a continuous system the slurry will then either overflow from one tank to the next, or be pumped to the next tank. Ultimately the “pregnant” solution is separated from the slurry using some form of liquid/solid separation process, and the solution passes on to the next phase of recovery.
In vat leaching the solids are loaded into the vat, once full the vat is flooded with a leaching solution. The solution drains from the tank, and is either recycled back into the vat or is pumped to the next step of the recovery process.
As mentioned previously tanks are equipped with agitators to keep the solids in suspension in the vats and improve the solid to liquid to gas contact. Agitation is further assisted by the use of tank baffles to increase the efficiency of agitation and prevent centrifuging of slurries in circular tanks.
Aside from chemical requirements several key factors influence extraction efficiency:
 Retention time - refers to the time spent in the leaching system by the solids. This is calculated as the total volumetric capacity of the leach tank/s divided by the volumetric throughput of the solid/liquid slurry. Retention time is commonly measured in hours for precious metals recovery. A sequence of leach tanks is referred to as a leach "train", and retention time is measured considering the total volume of the leach train. The desired retention time is determined during the testing phase, and the system is then designed to achieve this.
 Size - The ore must be ground to a size that exposes the desired mineral to the leaching agent (referred to as “liberation”), and in tank leaching this must be a size that can be suspended by the agitator. In vat leaching this is the size that is the most economically viable, where the recovery achieved as ore is ground finer is balanced against the increased cost of processing the material.
 Slurry density - The slurry density (percent solids) determines retention time. The settling rate and viscosity of the slurry are functions of the slurry density. The viscosity, in turn, controls the gas mass transfer and the leaching rate.
 Numbers of tanks - Agitated tank leach circuits are typically designed with no less than four tanks and preferably more to prevent short-circuiting of the slurry through the tanks.
 Dissolved gas - Gas is often injected below the agitator or into the vat to obtain the desired dissolved gas levels – typically oxygen, in some base metal plants sulphur dioxide may be required.
 Reagents - Adding and maintaining the appropriate amount of reagents throughout the leach circuit is critical to a successful operation. Adding insufficient quantities of reagents reduces the metal recovery but adding excess reagents increases the operating costs without recovering enough additional metal to cover the cost of the reagents.
The tank leaching method is commonly used to extract gold and silver from ore.
In-situ leach
In-situ leaching (ISL), also called in-situ recovery (ISR) or solution mining, is a leaching process used to recover minerals such as copper and uranium through boreholes drilled into a deposit, in situ.
The process initially involves drilling of holes into the ore deposit. Explosive or hydraulic fracturing may be used to create open pathways in the deposit for solution to penetrate. Leaching solution is pumped into the deposit where it makes contact with the ore. The solution bearing the dissolved ore content is then pumped to the surface and processed. This process allows the extraction of metals and salts from an ore body without the need for conventional mining involving drill-and-blast, open-cut or underground mining.
In-situ leach mining involves pumping of a leachate solution into the ore body via a borehole, which circulates through the porous rock dissolving the ore and is extracted via a second borehole.
The leachate solution varies according to the ore deposit: for salt deposits the leachate can be fresh water into which salts can readily dissolve. For copper, acids are generally needed to enhance solubility of the ore minerals within the solution. For uranium ores, the leachate may be acid or sodium bicarbonate.
Soluble salts
In-situ leaching is widely used to extract deposits of water-soluble salts such as sylvite (potash), halite (rock salt, sodium chloride), and sodium sulfate. It has been used in the US state of Colorado to extract nahcolite (sodium bicarbonate). In-situ leaching is often used when the deposits are too deep, or the beds too thin for conventional underground mining.
Uranium
Solutions used to dissolve uranium ore are either acid (sulfuric acid or less commonly nitric acid) or carbonate (sodium bicarbonate, ammonium carbonate, or dissolved carbon dioxide). Dissolved oxygen is sometimes added to the water to mobilize the uranium. ISL of uranium ores started in the United States and the Soviet Union in the early 1960s. The first uranium ISL in the US was in the Shirley Basin in the state of Wyoming, which operated from 1961-1970 using sulfuric acid. Since 1970, all commercial-scale ISL mines in the US have used carbonate solutions.
In-situ recovery involves the extraction of uranium-bearing water (grading as low as 0.05% U3O8). The extracted uranium solution is then filtered through resin beads. Through an ion exchange process, the resin beads attract uranium from the solution. Uranium loaded resins are then transported to a processing plant, where U3O8 is separated from the resin beads and yellowcake is produced. The resin beads can then be returned to the ion exchange facility where they are reused.
At the end of 2008 there were four in-situ leaching uranium mines operating in the United States, all using sodium bicarbonate. ISL produces 90% of the uranium mined in the US.
Significant ISL mines are operating in Kazakhstan and Australia. ISL mining accounted for 41% of the world's uranium production in 2010.
Copper
In-situ leaching of copper was done by the Chinese by 977 AD, and perhaps as early as 177 BC. Copper is usually leached using acid (sulfuric acid or hydrochloric acid), then recovered from solution by solvent extraction electrowinning (SX-EW) or by chemical precipitation.
Ores most amenable to leaching include the copper carbonates malachite and azurite, the oxide tenorite, and the silicate chrysocolla. Other copper minerals, such as the oxide cuprite and the sulfide chalcocite may require addition of oxidizing agents such as ferric sulfate and oxygen to the leachate before the minerals are dissolved. The ores with the highest sulfide contents, such as bornite and chalcopyrite will require more oxidants and will dissolve more slowly. Sometimes oxidation is speeded by the bacteria Thiobacillus ferrooxidans, which feeds on sulfide compounds.
Copper ISL is often done by stope leaching, in which broken low-grade ore is leached in a current or former conventional underground mine. The leaching may take place in backfilled stopes or caved areas. In 1994, stope leaching of copper was reported at 16 mines in the US. At the San Manuel mine in the US state of Arizona, ISL, underground mining, and open-pit mining were being done simultaneously in different parts of the same ore body.
Gold
In-situ leaching has not been used on a commercial scale for gold mining. A three-year pilot program was undertaken in the 1970s to in-situ leach gold ore at the Ajax mine in the Cripple Creek district in the US, using a chloride and iodide solution. After obtaining poor results, perhaps because of the complex telluride ore, the test was halted.
Environmental Concerns: ISL Mining of Uranium
In the USA legislation requires that the water quality in the affected aquifer be restored so as to enable its pre-mining use. Usually this is potable water or stock water (usually less than 500 ppm total dissolved solids), and while not all chemical characteristics can be returned to those pre-mining, the water must be usable for the same purposes as before. Often it needs to be treated by reverse osmosis, giving rise to a problem in disposing of the concentrated brine stream from this. The usual radiation safeguards are applied at an ISL Uranium mining operation, despite the fact that most of the ore body’s radioactivity remains well underground and there is hence minimal increase in radon release and no ore dust. Employees are monitored for alpha radiation contamination and personal dosimeters are worn to measure exposure to gamma radiation. Routine monitoring of air, dust and surface contamination are undertaken.
The leaching liquid used for in-situ leaching contains the leaching agent ammonium carbonate for example, or - particularly in Europe - sulfuric acid. This method can only be applied if the uranium deposit is located in porous rock, confined in impermeable rock layers.
The advantages of this technology are:
 Reduced hazards for the employees from accidents, dust, and radiation,
 Low cost, no need for large uranium mill tailings deposits.
The disadvantages of the in-situ leaching technology are:
 Risk of spreading of leaching liquid outside of the uranium deposit, involving subsequent groundwater contamination
 Unpredictable impact of the leaching liquid on the rock of the deposit, and
 the impossibility of restoring natural groundwater conditions after completion of the leaching operations.
Moreover, in-situ leaching releases considerable amounts of radon, and produces certain amounts of waste slurries and waste water during recovery of the uranium from the liquid.
After termination of an in-situ leaching operation, the waste slurries produced must be safely disposed, and the aquifer, contaminated from the leaching activities, must be restored. Groundwater restoration is a very tedious process that is not yet fully understood. So far, it is not possible to restore groundwater quality to previous conditions.
Newbery-Vautin chlorination process
The Newbery-Vautin chlorination process is a process to extract gold from its ore using chlorination developed by James Cosmo Newbery and Claude Vautin.
The process of extracting gold from ores by absorption of the precious metal in chlorine gas, from which it is reduced to a metallic state, is not a very new discovery. It was first introduced by Karl Friedrich Plattner around 1848, and at that time promised to revolutionize the processes for gold extraction. By degrees it was found that only a very clever chemist could work this process with practically perfect results, for many reasons. Lime and magnesia might be contained in the quartz, and would be attacked by the chlorine. These consume the reagents without producing any results, earthy particles would settle and surround the small gold and prevent chlorination, then lead and zinc or other metals in combination with the gold would also be absorbed by the chlorine; or, again, from some locally chemical peculiarity in the water or the ore, gold held in solution by the water might be again precipitated in the tailings before filtration was complete, and thus be lost. Henderson, Clark, De Lacy, Mears, and Deacon, all introduced improvements, or what were claimed to be improvements, on Plattner, but these chiefly failed because they did not cover every particular variety of case which gold extraction presented. Therefore, where delicate chemical operations were necessary for success, practice generally failed from want of knowledge on the part of the operator, and many times extensive plants have been pronounced useless from this cause alone. Hence it is not to be wondered that processes requiring such care and uncommon knowledge are not greatly in favor.
Claude Theodore James Vautin, a gentleman possessed of much practical experience of gold mining and extraction in Queensland, Australia, together with James Cosmo Newbery, analytical chemist to the government of Victoria, have developed a process which they claim to combine all the advantages of the foregoing methods, and by the addition of certain improvements in the machinery and mode of treatment to overcome the difficulties which have hitherto prevented the general adoption of the chlorination process.
The materials for treatment—crushed and roasted ore, or tailings, as the case may be—are put into the hopper above the revolving barrel, or chlorinator. This latter is made of iron, lined with wood and lead, and sufficiently strong to bear a pressure of 100 lb. to the square inch, its capacity being about 30 cwt. of ore. The charge falls from the hopper into the chlorinator. Water and chlorine-producing chemicals are added—generally sulphuric acid and chloride of lime—the manhole cover is replaced and screwed down so as to be gas tight. On the opposite side of the barrel there is a valve connected with an air pump, through which air to about the pressure of four atmospheres is pumped in, to liquefy the chlorine gas that is generated, after which the valve is screwed down. The barrel is then set revolving at about ten revolutions a minute, the power being transmitted by a friction wheel. According to the nature of the ore, or
the size of the grains of gold, this movement is continued from one to four hours, during which time the gold, from combination with the chlorine gas, has formed a soluble gold chloride, which has all been taken up by the water in the barrel. The chlorinator is then stopped, and the gas and compressed air allowed to escape from the valve through a rubber hose into a vat of lime water. This is to prevent the inhalation of any chlorine gas by the workmen. The manhole cover is now removed and the barrel again set revolving, by which means the contents are thrown automatically into the filter below. This filter is an iron vat lined with lead. It has a false bottom, to which is connected a pipe from a vacuum pump working intermittently. As soon as all the ore has fallen from the chlorinator into the filter, the pump is set going, a partial vacuum is produced in the chamber below the false bottom in the filter, and very rapid filtration results. By this means all the gold chlorides contained in the wet ore may be washed out, a continual stream being passed through it while filtration is going on. The solution running from the filter is continually tested, and when found free from gold, the stream of water is stopped, as is also the vacuum pump. The filter is then tipped up into a truck below, and the tailings run out to the waste heap. The process of washing and filtration occupies about an hour, during which time another charge may be in process of treatment in the chlorinator above. The discharge from the filter and the washings are run into a vat, and from this they are allowed to pass slowly through a tap into a charcoal filter. During the passage of the liquid through the charcoal filter, the chloride of gold is decomposed and the gold is deposited on the charcoal, which, when fully charged, is burnt, the ashes are fused with borax in a crucible, and the gold is obtained.
Gold cyanidation
Gold cyanidation (also known as the cyanide process or the MacArthur-Forrest process) is a metallurgical technique for extracting gold from low-grade ore by converting the gold to a water soluble coordination complex. It is the most commonly used process for gold extraction. Production of reagents for mineral processing to recover gold, copper, zinc and silver represents approximately 13% of cyanide consumption globally, with the remaining 87% of cyanide used in other industrial processes such as plastics, adhesives, and pesticides. Due to the highly poisonous nature of cyanide, the process is controversial and its usage is banned in a number of countries and territories.
In 1783 Carl Wilhelm Scheele discovered that gold dissolved in aqueous solutions of cyanide. He had earlier discovered cyanide salts. Cyanide was not applied to extraction of gold ores until 1887, when the MacArthur-Forrest Process was developed in Glasgow, Scotland by John Stewart MacArthur, funded by the brothers Dr Robert and Dr William Forrest.
The chemical reaction for the dissolution of gold, the "Elsner Equation", follows:
4 Au + 8 NaCN + O2 + 2 H2O → 4 Na[Au(CN)2] + 4 NaOH
In this redox process, oxygen removes, via a two-step reaction, one electron from each gold atom to form the complex Au(CN)2- ion.
The ore is comminuted using grinding machinery. Depending on the ore, it is sometimes further concentrated by froth flotation or by centrifugal (gravity) concentration. Water is added to produce a slurry or pulp. The alkaline ore slurry can be combined with a solution of sodium cyanide or potassium cyanide, however many operations utilize calcium cyanide, which is more cost effective.
To prevent the creation of toxic hydrogen cyanide during processing, lime (calcium hydroxide) or soda (sodium hydroxide) is added to the extracting solution to ensure that the acidity during cyanidation is maintained over pH 10.5 - strongly alkaline. Lead nitrate can improve gold leaching speed and quantity recovered, particularly in processing partially oxidized ores.
Oxygen is one of the reagents consumed during cyanidation, and a deficiency in dissolved oxygen slows leaching rate. Air or pure oxygen gas can be purged through the pulp to maximize the dissolved oxygen concentration. Intimate oxygen-pulp contactors are used to increase the partial pressure of the oxygen in contact with the solution, thus raising the dissolved oxygen concentration much higher than the saturation level at atmospheric pressure. Oxygen can also be added by dosing the pulp with hydrogen peroxide solution.
In some ores, particularly those that are partially sulfidized, aeration (prior to the introduction of cyanide) of the ore in water at high pH can render elements such as iron and sulfur less reactive to cyanide, and therefore the gold cyanidation process more efficient. Specifically, the oxidation of iron to iron (III) oxide and subsequent precipitation as iron hydroxide minimizes loss of cyanide from the formation of ferrous cyanide complexes. The oxidation of sulfur compounds to sulfate ions avoids the consumption of cyanide to thiocyanate (SCN-) byproduct.
In order of decreasing economic efficiency, the common processes for recovery of the solubilized gold from solution are (certain processes may be precluded from use by technical factors):
 Carbon in pulp
 Electrowinning
 Merrill-Crowe process
Cyanide remediation processes
The various species of cyanide that remain in tails streams from gold plants are potentially toxic, and on some operations the waste streams are processed through a detoxification process prior to tails deposition. This reduces the concentrations of these cyanide compounds,
but does not completely eliminate them from the stream. The two major processes utilised are the INCO licenced process or the Caro’s acid process. Both processes utilise oxidants to oxidise cyanide to cyanate, which is not as toxic as the cyanide ion, and which can then react to form carbonates and ammonia. The Inco process can typically reduce cyanide concentrations to below 50 mg/L, while the Caro’s acid process can reduce cyanide levels to between 10 and 50 mg/L, with the lower concentrations achievable in solution streams rather than slurries. Hydrogen peroxide and alkaline chlorination can also be used, although these are typically less common.
One of the alternative oxidants for the degradation of cyanides that has been attracting industrial interest is Caro’s acid – peroxomonosulphuric acid (H2SO5). Caro’s acid converts cyanide to cyanate. Cyanate then hydrolyses in the water to ammonium and carbonate ions. The Caro's acid process is able to achieve discharge levels of WAD below 50 mg/L, which is generally suitable for discharge to tailings. Generally, the best application of this process is with tailings slurries containing low to moderate initial levels of cyanide and when treated cyanide levels of less than about 10 to 50 mg/L are required.
Over 90 mines worldwide now use an Inco SO2/air detoxification circuit to convert cyanide to the much less toxic cyanate before waste is discharged to a tailings pond. Typically, this process blows compressed air through the tailings while adding sodium metabisulfite which releases SO2, lime to maintain the pH at around 8.5, and copper sulfate as a catalyst if there is insufficient copper in the ore extract. This procedure can reduce concentrations of "Weak Acid Dissociable" (WAD) cyanide to below the 10 ppm mandated by the EU's Mining Waste Directive. This level compares to the 66-81 ppm free cyanide and 500-1000 ppm total cyanide in the pond at Baia Mare. Remaining free cyanide degrades in the pond, while cyanate ions hydrolyse to ammonium. Recent studies show that residual cyanide trapped in the gold-mine tailings causes persistent release of toxic metals (e.g. mercury) into the groundwater and surface water systems.
Effects on the environment
Despite being used in 90% of gold production, gold cyanidation is controversial due to the toxic nature of cyanide. Although aqueous solutions of cyanide degrade rapidly in sunlight, the less-toxic products, such as cyanates and thiocyanates, may persist for some years. The famous disasters have killed few people — humans can be warned not to drink or go near polluted water — but cyanide spills can have a devastating effect on rivers, sometimes killing everything for several miles downstream. However, the cyanide is soon washed out of river systems and, as long as organisms can migrate from unpolluted areas upstream, affected areas can soon be repopulated. In the Someș River below Baia Mare, the plankton returned to 60% of normal within 16 days of the spill. Such spills have prompted fierce protests at new mines that involve
use of cyanide, such as Roşia Montană in Romania, Lake Cowal in Australia, Pascua Lama in Chile, and Bukit Koman in Malaysia.
2. Amalgamation
An amalgam is a substance formed by the reaction of mercury with another metal. Almost all metals can form amalgams with mercury, a notable exception being iron. Silver-mercury amalgams are important in dentistry, and gold-mercury amalgam is used in the extraction of gold from ore.
Mercury has been used in the gold and silver mining methods because of the convenience and ease with which mercury will amalgamate with them. In gold placer mining, in which minute specks of gold are washed from sand or gravel deposits, mercury was often used to separate the gold from other heavy minerals.
After all of the practical metal had been taken out from the ore, the mercury was dispensed down a long copper trough, which formed a thin coating of mercury on the exterior. The waste ore was then transferred down the trough, and any gold in the waste was amalgamated with the mercury. This coating would sometimes get scraped off and refined to get rid of the mercury, leaving behind somewhat high purity gold.
Mercury amalgamation was first useful to silver ores with the development of the patio process in Mexico in 1557. There were also additional amalgamation processes that were created for processing silver ores, including pan amalgamation and the Washoe process.
Gold extraction (mining)
This has proved effective where gold fines ("flour gold") would not be extractable from ore using hydro-mechanical methods. Large amounts of mercury were used in placer mining, where deposits composed largely of decomposed granite slurry were separated in long runs of "riffle boxes", with mercury dumped in at the head of the run. The amalgam formed is a heavy solid mass of dull gray color. (The use of mercury in 19th century placer mining in California, now prohibited, has caused extensive pollution problems in riverine and estuarine environments, ongoing to this day.) Sometimes substantial slugs of amalgam are found in downstream river and creek bottoms by amateur wet-suited miners seeking gold nuggets with the aid of an engine-powered water vacuum mounted on a float.
Where stamp mills were used to crush gold-bearing ore to fines, a part of the extraction process involved the use of mercury-wetted copper plates, over which the crushed fines were washed. A periodic scraping and re-mercurizing of the plate resulted in amalgam for further processing.
Amalgam obtained by either process was then heated in a distillation retort, recovering the mercury for reuse and leaving behind the gold. As this released mercury vapors to the atmosphere, the process could induce adverse health effects and long term pollution.
Today, mercury amalgamation has been replaced by other methods to recuperate gold and silver from ore in developed nations. Hazards of mercurial toxic waste have played a major role in the phasing out of the mercury amalgamation processes. However mercury amalgamation is still regularly used by small-scale gold placer miners (often illegally) & particularly in developing countries.
Mercury salts are, compared to mercury metal and amalgam, highly toxic due to their solubility in water. The presence of these salts in water can be detected with a probe that uses the readiness of mercury ions to form an amalgam with copper. A nitric acid solution of salts under investigation is applied to a piece of copper foil and any mercury ions present will leave spots of silvery-coloured amalgam. Silver ions leave similar spots but are easily washed away, making this a means of distinguishing silver from mercury.
The redox reaction involved where mercury oxidizes the copper is:
Hg2+ + Cu → Hg + Cu2+
Patio process
The patio process was a process used to extract silver from ore. The process was invented by Bartolomé de Medina in Pachuca, New Spain (Mexico), in 1554. The patio process was the first process to use mercury amalgamation to recover silver from ore. It replaced smelting as the primary method of extracting silver from ore at Spanish colonies in the Americas. Other amalgamation processes were later developed, importantly the pan amalgamation process, and its variant, the Washoe process. The silver separation process generally differed from gold parting and gold extraction, although amalgamation with mercury was also sometimes used to extract gold.
Basic elements of the patio process
Silver ores were crushed (typically either in "arrastras" or stamp mills) to a fine slime which was mixed with salt, water, magistral (essentially an impure form of copper sulfate), and mercury, and spread in a 1-to-2-foot-thick (0.30 to 0.61 m) layer in a patio, (a shallow-walled, open enclosure). Horses were driven around on the patio to further mix the ingredients. After weeks of mixing and soaking in the sun, a complex reaction converted the silver to native metal, which formed an amalgam with the mercury and was recovered. The amount of salt and copper sulfate varied from one-quarter to ten pounds of one or the other, or both, per ton of ore treated. The decision of how much of each ingredient to add, how much mixing was needed,
and when to halt the process depended on the skill of an azoguero (English: quicksilver man). The loss of mercury in amalgamation processes is generally one to two times the weight of silver recovered.
The patio process was the first form of amalgamation. However, it is unclear whether this process or a similar process—in which amalgamation occurred in heated vats rather than open patios—was the predominant form of amalgamation in New Spain, as the earliest known illustration of the patio process dates from 1761. There is substantial evidence that both processes were used from an early date in New Spain, while open patios were never adopted in Peru. Both processes required that ore be crushed and refiners quickly established mills to process ore once amalgamation was introduced. By the seventeenth century, water-powered mills became dominant in both New Spain and Peru.
Due to amalgamation's reliance upon mercury, an expansion of mercury production was central to the expansion of silver production. From shortly after the invention of mercury amalgamation to the end of the colonial period, the Spanish crown maintained a monopoly on mercury production and distribution, ensuring a steady supply of royal income. Fluctuations in mercury prices generally resulted in corresponding increases and decreases in silver production.
Pan amalgamation
The Pan amalgamation process is a method to extract silver from ore, using mercury. The process was widely used from 1609 through the 19th century; it is no longer used.
One drawback of the patio process was the long treatment time, usually weeks. Alvaro Alonso Barba invented the faster pan process (in Spanish the cazo or fondo process) in 1609 in Potosí, Bolivia, in which ore was mixed with salt and mercury (and sometimes copper(II) sulfate) and heated in shallow copper vessels. The treatment time was reduced to 10 to 20 hours. Whether patio or pan amalgamation was used at a particular location often depended on climate (warmer conditions speeded the patio process) and the availability and cost of fuel to heat the pans.
The amount of salt and copper(II) sulfate varied from one-quarter to ten pounds of one or the other, or both, per ton of ore treated. The loss of mercury in amalgamation processes was generally one to two times the weight of silver recovered.
Washoe process
The Washoe process, a variation of pan amalgamation, was developed in the 1860s by Almarin Paul and others, to work the ore from the Comstock Lode in Nevada, United States (Washoe
was an early name for the area). In the Washoe process, the copper pans were replaced by iron tanks with mechanical agitators. Each tank ("pan") was circular, and commonly held 1,200 to 1,500 pounds of ore that had been crushed to sand size. Water was added to make a pulp, and 60 to 70 pounds of mercury, along with one-half to three pounds each of salt (sodium chloride) and bluestone (copper(II) sulfate) were also added. A circular iron plate called a muller was mounted on a vertical shaft and lowered into the tank, and was rotated to provide both agitation and additional grinding. Heat was delivered to the tanks by steam pipes. The iron filings worn from the muller and pan proved to be an essential ingredient in the process.
Reese River process
A variation of the Washoe process was developed in the Reese River mining district around Austin, Nevada. The Washoe process was found not to work well for ores with arsenic or antimony sulfides, or with galena or sphalerite. In 1869, Carl A. Stetefeldt of Reno found that roasting the ore with salt converted the silver sulfides to silver chlorides, which could then be recovered in amalgamation pans. The process was introduced in the Reese River District in 1879, with great success.
ELECTROMETALLURGICAL ORE REDUCTION METHODS
Electrometallurgy is the field concerned with the processes of metal electrodeposition. There are four categories of these processes, but our focus here will be on electrowinning, the extraction of metal from ores.
1. Electrowinning
Electrowinning, also called electroextraction, is the electrodeposition of metals from their ores that have been put in solution or liquefied. Electrorefining uses a similar process to remove impurities from a metal. Both processes use electroplating on a large scale and are important techniques for the economical and straightforward purification of non-ferrous metals. The resulting metals are said to be electrowon.
In electrowinning, a current is passed from an inert anode through a liquid leach solution containing the metal so that the metal is extracted as it is deposited in an electroplating process onto the cathode. In electrorefining, the anodes consist of unrefined impure metal, and as the current passes through the acidic electrolyte the anodes are corroded into the solution so that the electroplating process deposits refined pure metal onto the cathodes.
Applications
The most common electrowon metals are lead, copper, gold, silver, zinc, aluminium, chromium, cobalt, manganese, and the rare-earth and alkali metals. For aluminium, this is the only production process employed. Several industrially important active metals (which react strongly with water) are produced commercially by electrolysis of their pyrochemical molten salts. Experiments using electrorefining to process spent nuclear fuel have been carried out. Electrorefining may be able to separate heavy metals such as plutonium, caesium, and strontium from the less-toxic bulk of uranium. Many electroextraction systems are also available to remove toxic (and sometimes valuable) metals from industrial waste streams.
Process
Most metals occur in nature in their oxidized form (ores) and thus must be reduced to their metallic forms. The ore is dissolved following some preprocessing in an aqueous electrolyte or in a molten salt and the resulting solution is electrolyzed. The metal is deposited on the cathode (either in solid or in liquid form), while the anodic reaction is usually oxygen evolution. Several metals are naturally present as metal sulfides; these include copper, lead, molybdenum, cadmium, nickel, silver, cobalt, and zinc. In addition, gold and platinum group metals are associated with sulfidic base metal ores. Most metal sulfides or their salts are electrically conductive and this allows electrochemical redox reactions to efficiently occur in the molten state or in aqueous solutions.
Fig. 3: Apparatus for electrolytic refining of copper
Some metals, including arsenic and nickel do not electrolyze out but remain in the electrolyte solution. These are then reduced by chemical reactions to refine the metal. Other metals, which during the processing of the target metal have been reduced but not deposited at the cathode, sink to the bottom of the electrolytic cell, where they form a substance referred to as anode sludge or anode slime. The metals in this sludge can be removed by standard pyrorefining methods.
Because metal deposition rates are related to available surface area, maintaining properly working cathodes is important. Two cathode types exist, flat-plate and reticulated cathodes, each with its own advantages. Flat-plate cathodes can be cleaned and reused, and plated metals recovered. Reticulated cathodes have a much higher deposition rate compared to flat-plate cathodes. However, they are not reusable and must be sent off for recycling. Alternatively, starter cathodes of pre-refined metal can be used, which become an integral part of the finished metal ready for rolling or further processing.
ORE REFINING OR PURIFICATION METHODS
The metal obtained after reduction process still contains some impurities which can be removed by applying the following methods.
PYROMETALLURGICAL METAL PURIFICATION METHODS
1. Cupellation
Cupellation is a refining process in metallurgy, where ores or alloyed metals are treated under high temperatures and controlled operations to separate noble metals, like gold and silver, from base metals like lead, copper, zinc, arsenic, antimony or bismuth, present in the ore. The process is based on the principle that precious metals do not oxidise or react chemically, unlike the base metals; so when they are heated at high temperatures, the precious metals remain apart and the others react forming slags or other compounds.
Since the Early Bronze Age, the process was used to obtain silver from smelted lead ores. By the Middle Ages and the Renaissance, cupellation was one of the most common processes for refining precious metals. By then, fire assays were used for assaying minerals, that is, testing fresh metals such as lead and recycled metals to know their purity for jewelry and coin making. Cupellation is still in use today.
Process
Large scale cupellation
Native silver is a rare element, although it exists as such. It is usually found in nature combined with other metals, or in minerals that contain silver compounds, generally in the form of sulfides such as galena (lead sulfide) or cerussite (lead carbonate). So the primary production of silver requires the smelting and then cupellation of argentiferous lead ores.
Lead melts at 327°C while silver melts at 960°C; when mixed, as in galena, the most common argentiferous lead ore, they have to be smelted at high temperatures in a reducing atmosphere
to produce argentiferous lead. The alloy is melted again at the high temperature of 900°C to 1000°C in a hearth or blast furnace, where air flow makes possible the oxidation of the lead. The lead oxidises to lead oxide (PbO) known as litharge, which captures the oxygen from the other metals present, while silver and gold remain on top of the liquid litharge. The latter is removed or absorbed by capillary action into the hearth linings. This chemical reaction may be viewed as:
(Ag+Cu) + Pb + O2 → (CuO+PbO) + Ag
The base of the hearth was dug in the form of a saucepan, and covered with an inert and porous material rich in calcium or magnesium such as shells, lime, or the ash from burning wood or bones. The lining had to be calcareous because lead reacts with silica (clay compounds) to form viscous lead silicate that prevents the needed absorption of litharge, whereas calcareous materials do not react with lead. Some of the litharge evaporates, and the rest is absorbed by the porous earth lining to form "litharge cakes".
Litharge cakes are usually circular or concavo-convex, about 15 cm in diameter. They are the most common archaeological evidence of cupellation in the Early Bronze Age. By their chemical composition, archaeologists can tell what kind of ore was treated, its main components, and the chemical conditions used in the process. This permits insights about production process, trade, social needs or economic situations.
Small scale cupellation
Small scale cupellation is based on the same principle as the one done in a cupellation hearth; the main difference lies in the amount of material to be tested or obtained. The minerals have to be crushed, roasted and smelted to concentrate the metallic components in order to separate the noble metals. By the Renaissance the use of the cupellation processes was diverse: assay of ores from the mines, testing the amount of silver in jewels or coins or for experimental purposes. It was carried out in small shallow recipients known as cupels.
As the main purpose of small scale cupellation was to assay and test minerals and metals, the material to be tested had to be carefully weighed. The assays were made in the cupellation or assay furnace, which needed to have windows and bellows to ascertain that the air oxidises the lead, as well as to be sure and prepared to take away the cupel when the process was over. Pure lead had to be added to the material being tested to guarantee the further separation of the impurities. After the litharge had been absorbed by the cupel, buttons of silver were formed and settled in the middle of the cupel. If the alloy also contained a certain amount of gold, it settled with the silver and both had to be separated by parting.
Cupels
The primary tool for small scale cupellation was the cupel. Cupels were manufactured in a very careful way. They used to be small vessels shaped in the form of an inverted truncated cone, made out of bone ashes. According to Georg Agricola, the best material was obtained from burned horns of deer although fish spines could work well. Ashes have to be ground into a fine and homogeneous powder and mixed with some sticky substance to mould the cupels. Moulds were made out of brass with no bottoms so that the cupels could be taken off. A shallow depression in the centre of the cupel was made with a rounded pestle. Cupel sizes depend on the amount of material to be assayed. This same shape has been maintained until the present.
Archaeological investigations as well as archaeometallurgical analysis and written texts from the Renaissance have demonstrated the existence of different materials for their manufacture; they could be made also with mixtures of bones and wood ashes, of poor quality, or moulded with a mixture of this kind in the bottom with an upper layer of bone ashes. Different recipes depend on the expertise of the assayer or on the special purpose for which it was made (assays for minting, jewelry, testing purity of recycled material or coins). Archaeological evidence shows that at the beginnings of small scale cupellation, clay cupels were used.
2. Parkes process
The Parkes process is a pyrometallurgical industrial process for removing silver from lead, during the production of bullion. It is an example of liquid-liquid extraction.
The process takes advantage of two liquid-state properties of zinc. The first is that zinc is immiscible with lead, and the other is that silver is 3000 times more soluble in zinc than it is in lead. When zinc is added to liquid lead that contains silver as a contaminant, the silver preferentially migrates into the zinc. Because the zinc is immiscible in the lead it remains in a separate layer and is easily removed. The zinc-silver solution is then heated until the zinc vaporizes, leaving nearly pure silver. If gold is present in the liquid lead, it can also be removed and isolated by the same process.
3. Polling
Polling is a method employed in the purification of copper which contains cuprous oxide as an impurity. The impure metal is melted and green wooden poles are used to agitate the molten impure copper. The heat of the copper makes the pole emit a gas which reduces the cuprous oxide to copper.
4. Distillation method
Volatile metals (Hg, Zn, Cd) are easily purified by distillation. The impure metals are heated in a retort and vapours of volatile metals are collected and condensed in a receiver leaving behind non-volatile impurities in the retort.
5. Zone refining of fractional crystallization
This method is employed to get metals of very high purity (Ge, Si, B, Ga, In). The method is based on the difference in solubility of impurities of molten and solid state of the metal. A movable heater is allowed to move across the impure metals rod from one end to the other end. The pure metal crystallises while the impurities pass on to the adjacent melted zone.
6. Mond’s Process
This method is employed for purification of nickel. Impure nickel is converted into volatile nickel carbonyl by reaction of CO at 60-80oC. Nickel carbonyl decomposes at 180oC to form pure nickel and CO.
Ni + 4CO Ni(CO)4 Ni +4CO
HYDROMETALLURGICAL PURIFICATION METHODS
1. Leaching
Bayer process
The Bayer process is the principal industrial means of refining bauxite to produce alumina (aluminum oxide). Bauxite, the most important ore of aluminum, contains only 30–54% aluminum oxide, (alumina), Al2O3, the rest being a mixture of silica, various iron oxides, and titanium dioxide. The aluminum oxide must be purified before it can be refined to aluminum metal.
Process
In the Bayer process, bauxite is digested by washing with a hot solution of sodium hydroxide, NaOH, at 175°C. This converts the aluminum oxide in the ore to sodium aluminate, 2NaAl(OH)4, according to the chemical equation:
Al2O3 + 2NaOH + 3H2O → 2NaAl(OH)4
The other components of bauxite do not dissolve. The solution is clarified by filtering off the solid impurities. The mixture of solid impurities is called red mud, and presents a disposal problem. Next, the alkaline solution is cooled, and aluminum hydroxide precipitates as a white, fluffy solid:
NaAl(OH)4 → Al(OH)3 + NaOH
Then, when heated to 980°C (calcined), the aluminum hydroxide decomposes to aluminum oxide, giving off water vapor in the process:
2Al(OH)3 → Al2O3 + 3H2O
A large amount of the aluminum oxide so produced is then subsequently smelted in the Hall–Héroult process in order to produce aluminum.
ELECTOMETALLURGICAL PURIFICATION METHODS
1. Electrolysis
The impure metal is made the anode while a thin sheet of pure metal acts as the cathode. The electrolytic solution consists of an aqueous solution of salt or a complex of the metal. On passing the current, pure metal is deposited on the cathode and an equivalent amount of the
metal gets dissolved from anode. The soluble impurities pass into the solution and the insoluble impurities collect below the anode as anode mud.
The Hall–Héroult process
The Hall–Héroult process is the major industrial process for the production of aluminum. It involves dissolving alumina in molten cryolite, and electrolyzing the molten salt bath to obtain pure aluminum metal.
Fig. 5: A Hall–Héroult industrial cell
Process
Aluminum cannot be produced by the electrolysis of an aluminum salt dissolved in water because of its high reactivity: the water (especially, hydronium ions which are its natural constituent) readily oxidizes elemental aluminum. The reduction of Al3+ is done by electrolysis of a molten aluminum salt. This is a water-free medium.
Alumina, Al2O3, is dissolved in an industrial carbon-lined vat of molten cryolite, Na3AlF6 (sodium hexafluoroaluminate), called a "cell". Aluminum oxide has a melting point of over 2,000 °C while pure cryolite has a melting point of 1,012 °C. With a small percentage of alumina
dissolved in it, cryolite has a melting point of about 1,000 °C. Some aluminum fluoride, AlF3 is also added into the process to reduce the melting point of the cryolite-alumina mixture.
The molten mixture of cryolite, alumina, and aluminum fluoride is then electrolyzed by passing a direct electric current through it. The electrochemical reaction causes liquid aluminum metal to be deposited at the cathode as a precipitate, while the oxygen from the alumina combines with carbon from the anode to produce carbon dioxide, CO2. An electric potential of three to five volts is needed to drive the reaction, and the rate of production is proportional to the electric current. An industrial-scale smelter typically consumes hundreds of thousands of amperes for each cell.
The oxidation of the carbon anode reduces the required voltage across each cell, increasing the electrical efficiency, at a cost of continually replacing the carbon electrodes with new ones, and also the cost of releasing carbon dioxide into the atmosphere. Hundreds of Hall–Héroult cells are usually connected electrically in series, and they are supplied with direct current (DC) from a single set of rectifiers that convert alternating current (AC) supplied to the factory into direct current. The very high electric current is supplied to the cells through heavy, low electrical resistance metal busbars made of pure aluminum or copper. The cells are electrically heated to reach the operating temperature with this current, and the anode regulator system varies the current passing through the cell by raising or lowering the anodes and changing the cell's resistance. If needed any cell can be bypassed by shunt busbars.
The liquid aluminum is taken out with the help of a siphon operating with a vacuum, in order to avoid having to use extremely high temperature valves and pumps. The liquid aluminum then may be transferred in batches or via a continuous hot flow line to a location where it is cast into aluminum ingots. The aluminum can either be cast into the form of final cast-aluminum products, or the ingots can be sent elsewhere such as a rolling mill to be pressed into sheets, or a wire-drawing mill producing aluminum wires and cables.
While solid cryolite is denser than solid aluminum at room temperature, the liquid aluminum product is denser than the molten cryolite at temperatures around 1,000 °C, and the aluminum sinks to the bottom of the electrolytic cell, where it is periodically collected. The tops and sides of the cells are covered with layers of solid cryolite which also act as thermal insulation. The unavoidable electric resistance within each cell produces sufficient heat to keep the cryolite-alumina mixture molten.
With the percentage of aluminum dissolved in each cell being depleted by the electrolysis in the molten cryolite, additional alumina is continually dropped into the cells to maintain the required level of alumina. Whenever a solid crust forms across the surface of the molten
cryolite-alumina, this crust is broken from time to time to allow the added alumina to fall into the molten cryolite and dissolve there.
The electrolysis process produces exhaust which escapes into the fume hood and is evacuated. The exhaust is primarily CO2 produced from the anode consumption and hydrogen fluoride (HF) from the cryolite and flux. HF is a highly corrosive and toxic gas, even etching glass surfaces. The gases are either treated or vented into the atmosphere; the former involving neutralization of the HF to its sodium salt, sodium fluoride. The particulates are also captured and reused using electrostatic or bag filters. The remaining CO2 is usually vented into the atmosphere.
The very large electric current passing through the electrolytic cells generates a powerful magnetic field, and this can stir the molten aluminum with magneto-hydrodynamic forces in properly-designed cells. The stirring of the molten aluminum in each cell typically increases its performance, but the purity of the aluminum is reduced, since it gets mixed with small amounts of cryolite and aluminum fluoride. If the cells are designed for no stirring, they can be operated with static pools of molten aluminum so that the impurities either rise to the top of the metallic aluminum, or else sink to the bottom, leaving high-purity aluminum in the middle.
Aluminum smelters are usually sited where inexpensive hydroelectric power is available. For some European smelters, the electric power produced by hydroelectric plants in countries such as Norway, Switzerland, and Austria is transmitted by high-voltage power lines to such places as Denmark, Sweden, Germany, and Italy to be used by aluminum and magnesium factories. Since aluminum factories require nearly-uniform supplies of electric current, they make the most of nearly-constant supplies of electric power, and these are also available close to many hydroelectric power plants.
An alternate source of power, used by Icelandic smelters, is geothermal electricity, which Iceland has in abundance owing to its location on the Mid-Atlantic Ridge, but cannot use domestically owing to its low population. Iceland thus imports raw aluminum ore and uses the Hall–Héroult process as a means of exporting its electricity and thus efficiently exploiting the abundance of geothermal energy.
Impact
Aluminum is the most abundant metallic element on Earth but it is rarely found in its elemental state. It occurs in many minerals but its primary commercial source is bauxite, a mixture of hydrated aluminum oxides and compounds of other elements such as iron. It is converted to aluminum oxide, Al2O3 by the Bayer process and then to metallic aluminum by the Hall–Héroult process.
Prior to the Hall–Héroult process, elemental aluminum was made by heating ore along with elemental sodium or potassium in a vacuum. The method was complicated and consumed materials that were in themselves expensive at that time. This meant the cost to produce a small amount of aluminum in the early 19th century was very high.
Early aluminum was more costly than gold or platinum. Bars of aluminum were exhibited alongside the French crown jewels at the Exposition Universelle of 1855, and Emperor Napoleon III of France was said to have reserved his few sets of aluminum dinner plates and eating utensils for his most honored guests.
Production costs using older methods did come down, but when aluminum was selected as the material for the cap/lightning rod to sit atop the Washington Monument in Washington, D.C., it was still more expensive than silver.
New production based on the Hall–Héroult process, in combination with cheaper electric power, helped make aluminum (and incidentally magnesium) an inexpensive commodity.
This in turn helped make it possible for pioneers like Hugo Junkers to utilize aluminum and aluminum-magnesium alloys to make items like metal airplanes by the thousands, or Howard Lund to make aluminum fishing boats.
Castner process
The Castner process is a process for manufacturing sodium metal by electrolysis of molten sodium hydroxide at approximately 330°C. Below that temperature the melt would solidify, above that temperature, the metal would start to dissolve in the melt.
Fig. 6: Diagram of Castner process apparatus
Process
The diagram shows a ceramic crucible with a steel cylinder suspended within. Both cathode (C) and anode (A) are made of iron or nickel. The temperature is cooler at the bottom and hotter at the top so that the sodium hydroxide is solid in the neck (B) and liquid in the body of the vessel. Sodium metal forms at the cathode but is less dense than the fused sodium hydroxide electrolyte. Wire gauze (G) confines the sodium metal to accumulating at the top of the collection device (P). The cathode reaction is
2Na+ + 2e– → 2Na
The anode reaction is
2OH– → ½O2 + H2O + 2e–
Despite the elevated temperature some of the water produced remains dissolved in the electrolyte. This water diffuses throughout the electrolyte and results in the reverse reaction taking place on the electrolyzed sodium metal:
Na + H2O → ½H2 + Na+ + OH–
with the hydrogen gas also accumulating at (P). This, of course, reduces the efficiency of the process.
Van Arkel process
This method is employed to obtain ultra-pure metals. The impure metal is converted into a volatile compound while the impurities are not affected. The volatile compound is then decomposed electrically to get the pure metal. Other metals that can be purified this method are Zr, V, W, Hf etc.

Notes On Chemical Mehods of Mineral Extraction

EXTRACTIVE METALLURGY II
INTRODUCTION
Extractive metallurgy is the process of extracting valuable metals from their natural sources (ore) and refining the extracted raw metals into a purer form for practical use. The field is an applied science, covering all aspects of the physical and chemical processes used to produce mineral-containing and metallic materials, sometimes for direct use as a finished product, but more often in a form that requires further physical processing which is generally the subject of physical metallurgy, ceramics and other disciplines within the broad field of materials science.
The field of extractive metallurgy encompasses many specialty sub-disciplines, each concerned with various physical and chemical processes that are steps in an overall process of producing a particular material. These specialties are generically grouped into the categories of mineral processing, hydrometallurgy, pyrometallurgy, and electrometallurgy. The distinction among these groups of sub-disciplines within extractive metallurgy is far from clear, and many commercially important metallurgical processes involve considerable overlap.
The theoretical basis of extractive metallurgy is underpinned by the more general sciences of physics, chemistry, and geology. Additionally, the practice of extractive metallurgy nearly always involves contributions from other scientific fields such as analytical chemistry and mineralogy.
Extractive metallurgists are interested in three primary streams: feed, concentrate (valuable metal oxide/sulfide), and tailings (waste). After mining, large pieces of the ore feed are broken through crushing and/or grinding in order to obtain particles small enough where each particle is either mostly valuable or mostly waste. Concentrating the particles of value in a form supporting separation enables the desired metal to be removed from waste products.
Mining may not be necessary if the ore body and physical environment are conducive to leaching. Leaching dissolves minerals in an ore body and results in an enriched solution. The solution is collected and processed to extract valuable metals.
Ore bodies often contain more than one valuable metal. Tailings of a previous process may be used as a feed in another process to extract a secondary product from the original ore. Additionally, a concentrate may contain more than one valuable metal. That concentrate would then be processed to separate the valuable metals into individual constituents.
PRINCIPLES OF EXTRACTION
The extraction of pure metal from its ore involves several physical and chemical methods. The suitability of the method depends upon the nature of ore, the properties of metals and local conditions. The extractive metallurgy of a metal involves the following operations:
1. Crushing of ores
2. Concentration of ores
3. Reduction
4. Purification of ores
ORE REDUCTION METHODS
Some of the methods commonly used to get free metal from ore are given below:
PYROMETALLURGICAL ORE REDUCTION METHODS
1. Smelting
Smelting is the most common pyrometallurgical process and involves the application of heat and a chemical reducing agent to decompose ore, driving off other elements as gasses or slag and leaving just the metal behind. The reducing agent is commonly a source of carbon such as coke. The carbon removes oxygen from the ore as carbon monoxide, leaving behind elemental metal. Other reducing agents in use today include H2, CO, water, Na, K, Mg and Al.
Iron Smelting:
Zone 1 (<950°C), reduction of Fe2O3, Fe3O4 takes place:
3Fe2O3(s) + CO(g) → 2Fe3O4(s) + CO2(g)
Fe3O4(s) + CO(g) → 3FeO(s) + CO2(g)
Zone 2 (950-1,000°C), chemically reserve zone, FeO is in equilibrium with the gaseous phase:
FeO(s) + CO(g) ↔ Fe(l) + CO2(g)
Zone 3 (950-1,050°C), the reduction of FeO by rising CO gas takes place:
FeO(s) + CO(g) → Fe(l) + CO2(g)
Zone 4 (>1,000-1,050°C), direct reduction of FeO by C takes place:
FeO(s) + C(s) → Fe(l) + CO(g)
Raceway zone (below Zone 4) (1,050°C), conversion of CO2 to CO:
C(s) + O2(g) → CO2(g)
CO2(g) + C(s) → 2CO(g) (Boudward reaction)
2C(s) + O2(g) ↔ 2CO(g) (Overall reaction)
The gangue consists of SiO2, Al2O3, P, S bearing minerals. Their removal proceeds thus:
CaO(s) + SiO2(s) → CaSiO3(l) (1,200°C)------slag
CaO(s) + P2O5(s) → Ca3(PO4)2(l) (1,200°C)--slag
Fig. 1: Reaction zones in a blast furnace
Flash smelting
Flash smelting is a smelting process for sulfur-containing ores including chalcopyrite. The process was developed in 1949 in Finland for smelting copper ore. It has also been adapted for nickel and lead production. The process uses the autogenic principle by using the energy contained in the sulfur and iron for melting the ore. In the process dried and powdered ore is discharged from a nozzle into a fluidized bed reactor fed with oxygen. The reduced metal melts, and drops to the bottom of a settling chamber. The flotation produces a large effective surface area of fine-grained concentrate particles. The process makes smelting more energy efficient and environmentally friendly. Sulfur is released mainly in its solid form, thus reducing atmospheric pollution. The process is today used for 50% of the world’s primary copper production. The other 50% is mainly produced from oxide ores, where the process cannot be applied.
Copper smelting
Ores of Cu
chalcopyrite (CuFeS2), chalcocite (Cu2S), covellite (CuS), bornite (2Cu2S·CuS·FeS), tetrahedrite (Cu3SbS3+x(Fe,Zn)6Sb2S9), malachite (CuCO3·Cu(OH)2), azurite (2CuCO3·Cu(OH)2), cuprite (Cu2O), chrysocolia (CuO·SiO2·2H2O) and tennantite (Cu12As4S13).
Roasting
The copper ore concentrate is placed in a roaster and is partially oxidized to produce “calcine” and sulphur dioxide gas:
2CuFeS2(s) + 3O2(g) → 2FeO(s) + 2CuS(s) + 2SO2(g)
Roasting is no longer in common use as direct smelting is now favoured.
Smelting
The calcine is then mixed with silica and coke and smelted in an exothermic reaction at 1200°C (above the melting point of copper, but below that of iron and silica) to form a liquid called “copper matte”. The high temperature allows reactions to proceed rapidly, and allow the matte and slag to melt, so they can be tapped out of the furnace. In copper recycling, this is the point where scrap copper is introduced. Several reactions occur:
FeO(s) + SiO2(s) → FeSiO3(l)
In a parallel reaction, iron sulfide is converted to slag:
2FeS(s) + 3O2 + SiO2(s) → 2FeSiO3(l) + 2SO2(g)
The slag is discarded or reprocessed to recover any remaining copper. The sulphuric acid is captured for use in earlier leaching processes.
Conversion to blister
The matte, which is produced in the smelter, contains around 70% copper primarily as copper sulfide as well as iron sulfide. The sulphur is removed at high temperature as sulphur dioxide by blowing air through molten matte:
2CuS + 3O2 → 2CuO + 2SO2
CuS + O2 → Cu + SO2
In a parallel reaction the iron sulfide is converted to slag:
2FeS + 3O2 → 2FeO + 2SO2
2FeO + 2SiO2 → 2FeSiO3
The purity of this product is 98%, it is known as blister because of the broken surface created by the escape of sulfur dioxide gas as the copper pigs or ingots are cast. By-products generated in the process are sulfur dioxide and slag. The sulfur dioxide is captured for use in earlier leaching processes.
Reduction
The blistered copper is put into an anode furnace (a furnace that uses the blister copper as anode) to get rid of most of the remaining oxygen. This is done by blowing natural gas through the molten copper oxide. When this flame burns green, indicating the copper oxidation spectrum, the oxygen has mostly been burned off. This creates copper at about 99% pure. The anodes produced from this are fed to the electrorefinery.
2. Self-reduction process
The sulphide ores of less electropositive metals like Hg, Pb, Cu, etc. are heated in air to convert part of sulphide ore into oxide, which then reacts with the remaining sulphide to give the metal and sulphur dioxide.
2HgS + 3O2 → 2HgO + 2SO2
2HgO + HgS → 3Hg + SO2
3. Roasting
Roasting is a step of the processing of certain ores. It is a pyrometallurgical process involving gas–solid reactions at elevated temperatures with the goal of purifying the metal component(s). Often before roasting, the ore has already been partially purified, e.g. by froth floatation. The concentrate is mixed with other materials to facilitate the process. The technology is useful but is also a serious source of air pollution.
Roasting consists of thermal gas–solid reactions, which can include oxidation, reduction, chlorination, sulfation, and pyrohydrolysis. In roasting, the ore or ore concentrate is treated with very hot air. This process is generally applied to sulphide minerals. During roasting, the sulfide is converted to an oxide, and sulfur is released as sulfur dioxide, a gas. For the ores Cu2S (chalcocite) and ZnS (sphalerite), balanced equations for the roasting are:
2Cu2S + 3O2 → 2Cu2O + 2SO2
2ZnS + 3O2 → 2ZnO + 2SO2
The gaseous product of sulfide roasting, sulfur dioxide (SO2) is often used to produce sulfuric acid. Many sulfide minerals contain other components such as arsenic that are released into the environment.
Up until the early 20th century, roasting was started by burning wood on top of ore. This would raise the temperature of the ore to the point where its sulfur content would become its source
of fuel, and the roasting process could continue without external fuel sources. Early sulfide roasting, was practiced in this manner, in "open hearth" roasters, which were manually stirred (a practice referred to as "rabbling") using rake-like tools to expose unroasted ore to oxygen as the reaction proceeded.
4. Calcination
Calcination (also referred to as calcining) is a thermal treatment process in the absence of air applied to ores and other solid materials to bring about a thermal decomposition, phase transition, or removal of a volatile fraction. The calcination process normally takes place at temperatures below the melting point of the product materials. Calcination is not the same process as roasting. In roasting, more complex gas–solid reactions take place between the furnace atmosphere and the solids.
The process of calcination derives its name from the Latin calcinare (to burn lime) due to its most common application, the decomposition of calcium carbonate (limestone) to calcium oxide (lime) and carbon dioxide (Fig. 2), in order to produce cement. The product of calcination is usually referred to in general as "calcine," regardless of the actual minerals undergoing thermal treatment. Calcination is carried out in furnaces or reactors (sometimes referred to as kilns or calciners) of various designs including shaft furnaces, rotary kilns, multiple hearth furnaces, and fluidized bed reactors.
Fig. 2: An oven for calcination of limestone
Examples of calcination processes include the following:
 decomposition of carbonate minerals, as in the calcination of limestone to drive off carbon dioxide;
 decomposition of hydrated minerals, as in the calcination of bauxite and gypsum, to remove crystalline water as water vapor;
 decomposition of volatile matter contained in raw petroleum coke;
 heat treatment to effect phase transformations, as in conversion of anatase to rutile or devitrification of glass materials
 removal of ammonium ions in the synthesis of zeolites
Calcination reactions usually take place at or above the thermal decomposition temperature (for decomposition and volatilization reactions) or the transition temperature (for phase transitions). This temperature is usually defined as the temperature at which the standard Gibbs free energy for a particular calcination reaction is equal to zero. For example, in limestone calcination, a decomposition process, the chemical reaction is
CaCO3 → CaO + CO2(g)
The standard Gibbs free energy of reaction is approximated as ΔG°r = 177,100 − 158 T (J/mol). The standard free energy of reaction is zero in this case when the temperature, T, is equal to 1121 K, or 848 °C.
In some cases, calcination of a metal may results in oxidation of the metal, leading to it gaining weight, e.g. Pb and Sn.
5. Liquation
Liquation is a metallurgical method for separating metals from an ore or alloy. The material must be heated until one of the metals starts to melt and drain away from the other and can be collected. This method was largely used to remove lead containing silver from copper, but it can also be used to remove antimony minerals from ore, and to refine tin.
Separating copper and silver
The first known use of Liquation on a large scale was in Germany in the mid-15th century. Liquation requires that the silver-rich copper first be melted with approximately three times its weight in lead, as silver has a greater affinity with lead most of the silver would end up within this rather than the copper. If the copper is assayed and found to contain too little silver for
liquation to be financially viable (around 0.31% is the minimum required) it is melted and allowed to settle so that much of the silver sinks towards the bottom. The ‘tops’ are then drawn off and used to produce copper while the silver-rich ‘bottoms’ are used in the liquation process. The copper-lead alloy created can be tapped off and cast into large plano-convex ingots known as ‘liquation cakes’. As the metals cool and solidify the copper and the silver-containing lead separate as they are immiscible with each other.
The ratio of lead to copper in these cakes is an important factor for the process to work efficiently. Three parts copper to 8-12 parts lead is recommended. The copper must be assayed to accurately determine how much silver it contains, for copper rich in silver the top end of this ratio was used to make sure the maximum amount of silver possible would end up within the lead. However there also needs to be enough copper to allow the cakes to keep their shape once most of the lead has drained away, too much copper and it would trap some of the lead within and the process would be very inefficient.
The size of these cakes remained consistent from 1556 to the 19th century when the process became obsolete. They were usually between 2½ to 3½ inches (6.4 to 8.9 cm) thick, about 2 feet (0.61m) in diameter and weighed from 225 to 375 lbs (102 kg to 170 kg). This consistency is not without reason as the size of the cakes is very important to the smooth running of the liquation process. If the cakes are too small, the product would not be worth the time and costs spent on the process, if they are too large then the copper would begin to melt before the maximum amount of lead has drained away.
The cakes are heated in a liquation furnace, usually four or five at once, to a temperature above the melting point of lead (327°C), but below that of copper (1084°C), so that the silver-rich lead melts and flows away. As the melting point of lead is so low a high temperature furnace is not required and it can be fuelled with wood. It is important that this takes place in a reducing atmosphere, i.e. one with little oxygen, to avoid the lead oxidising, the cakes are therefore well covered by charcoal and little air is allowed into the furnace. It is impossible to stop some of the lead oxidising however and this drops down and forms spiky projections known as ‘liquation thorns’ in the channel underneath the hearth.
The older and relatively simple method of cupellation can then be used to separate the silver from the lead. If the lead is assayed and found not to contain enough silver to make the cupellation process worthwhile it is reused in liquation cakes until it has sufficient silver.
The ‘exhausted liquation cakes’ which still contain some lead and silver are ‘dried’ in a special furnace which is heated to a higher temperature under oxidising conditions. This is essentially just another stage of liquation and most of the remaining lead is expelled and oxidised to form
liquation thorns, though some remains as lead metal. The copper can then be refined to remove other impurities and produce copper metal.
Waste products can be reused to produce new liquation cakes to try and minimise loss of metals, especially silver. The waste products are mostly in the form of liquation thorns from the liquation and the drying process but there are also some slags produced.
HYDROMETALLURGICAL ORE REDUCTION METHODS
1. Leaching
Leaching is a widely used extractive metallurgy technique which converts metals into soluble salts in aqueous media. Compared to pyrometallurgical operations, leaching is easier to perform and much less harmful, because no gaseous pollution occurs. Drawbacks of leaching are the highly acidic and in some cases toxic residual effluent, and its lower efficiency caused by the low temperatures of the operation, which dramatically affect chemical reaction rates.
There are a variety of leaching processes, usually classified by the types of reagents used in the operation. The reagents required depend on the ores or pretreated material to be processed. A typical feed for leaching is either oxide or sulfide.
For material in oxide form, a simple acid leaching reaction can be illustrated by the zinc oxide leaching reaction:
ZnO + H2SO4 → ZnSO4 + H2O
In this reaction solid ZnO dissolves, forming soluble zinc sulfate.
In many cases other reagents are used to leach oxides. For example, in the metallurgy of aluminium, aluminium oxide is subject to leaching by alkali solutions:
Al2O3 + 3H2O + 2NaOH → 2NaAl(OH)4
Leaching of sulfides is a more complex process due to the refractory nature of sulfide ores. It often involves the use of pressurized vessels, called autoclaves. A good example of the autoclave leach process can be found in the metallurgy of zinc. It is best described by the following chemical reaction:
2ZnS + O2 + 2H2SO4 → 2ZnSO4 + 2H2O + 2S
This reaction proceeds at temperatures above the boiling point of water, thus creating a vapor pressure inside the vessel. Oxygen is injected under pressure, making the total pressure in the autoclave more than 0.6 MPa.
The leaching of precious metals such as gold can be carried out with cyanide or ozone under mild conditions. A lixiviant is a liquid medium used in hydrometallurgy to selectively extract the desired metal from the ore or mineral. It assists in rapid and complete leaching. The metal can be recovered from it in a concentrated form after leaching. Lixiviant in a solution may be acidic or basic in nature. The most commonly used acid is H2SO4.
Heap leaching
Heap leaching is an industrial mining process to extract precious metals, copper, uranium, and other compounds from ore via a series of chemical reactions that absorbs specific minerals and then re-separate them after their division from other earth materials. Comparable to in situ mining, heap leach mining differs in that it uses a liner to place amounts of ore on, then adds the chemicals via drip systems to the ore, whereas in situ leaching lacks these pads and pulls pregnant solution up to obtain the minerals. This method is only slightly friendlier environmentally, however, and has still seen copious amounts of negative feedback from both environmentalists and health experts in the past twenty or more years. Since its original peak of popularity in the 1970s, the heap leach technique has been applied throughout the earth, but due to recent increases in negative environmental impact assessments has received more discussion regarding rehabilitation than perpetuation of these types of techniques. Yet this method continues to be a profit-earning endeavor for many mining companies across the globe.
The process has ancient origins; one of the classical methods for the manufacture of copperas (iron sulfate) was to heap up iron pyrite and collect the leachate from the heap, which was then boiled with iron to produce iron sulfate
The mined ore is usually crushed into small chunks and heaped on an impermeable plastic and/or clay lined leach pad where it can be irrigated with a leach solution to dissolve the valuable metals. While sprinklers are occasionally used for irrigation, more often operations use drip irrigation to minimize evaporation, provide more uniform distribution of the leach solution, and avoid damaging the exposed mineral. The solution then percolates through the heap and leaches both the target and other minerals. This process, called the "leach cycle," generally takes from one or two months for simple oxide ores (e.g., most gold ores) to two years (for nickel laterite ores). The leach solution containing the dissolved minerals is then collected, treated in a process plant to recover the target mineral and in some cases precipitate other minerals, and then recycled to the heap after reagent levels are adjusted. Ultimate recovery of the target mineral can range from 30% of contained (run-of-mine dump leaching sulfide copper ores) to over 90% for the easiest to leach ores (some oxide gold ores).
In recent years, the addition of an agglomeration drum has improved on the heap leaching process by allowing for a more efficient leach. The rotary drum agglomerator works by taking the crushed ore fines and agglomerating them into more uniform particles. This makes it much easier for the leaching solution to percolate through the pile, making its way through the channels between particles.
The addition of an agglomeration drum also has the added benefit of being able to pre-mix the leaching solution with the ore fines, to achieve a more concentrated, homogeneous mixture, and allowing the leach to begin prior to the heap.
Precious metals
The crushed ore is irrigated with a dilute alkaline cyanide solution. The solution containing the dissolved precious metals ("pregnant solution") continues percolating through the crushed ore until it reaches the liner at the bottom of the heap where it drains into a storage (pregnant solution) pond. After separating the precious metals from the pregnant solution, the dilute cyanide solution (now called "barren solution") is normally re-used in the heap-leach-process or occasionally sent to an industrial water treatment facility where the residual cyanide is treated and residual metals are removed. In very high rainfall areas, such as the tropics, in some cases there is surplus water that is then discharged to the environment, after treatment, posing possible water pollution if treatment is not properly carried out.
The production of one gold ring through this method can generate 20 tons of waste material.
During the extraction phase, the gold ions form complex ions with the cyanide:
Au+(s) + 2CNˉ(aq) → Au(CN)2ˉ(aq)
Recuperation of the gold is readily achieved with a redox-reaction:
2Au(CN)2ˉ(aq) + Zn(s) → Zn(CN)4ˉ(aq) + 2Au(s)
The most common methods to remove the gold from solution are either using activated carbon to selectively absorb it, or the Merrill-Crowe process where zinc powder is added to cause a precipitation of gold and zinc. The fine product can be either doré (gold-silver bars) or zinc-gold sludge that is then refined elsewhere.
Copper Ores
The method is similar to the cyanide method, above, except sulfuric acid is used to dissolve copper from its ores. The acid is recycled from the solvent extraction circuit and reused on the leach pad. A byproduct is iron (II) sulfate, jarosite, which is produced as a byproduct of leaching pyrite, and sometimes even the same sulfuric acid that is needed for the process. Both oxide
and sulfide ores can be leached, though the leach cycles are much different and sulfide leaching requires a bacterial or "bio-leach" component. The largest copper heap leach operations are in Chile, Peru, and the southwestern United States.
Although the heap leaching is a low cost-process, it normally has recovery rates of 60-70%, although there are exceptions. It is normally most profitable with low-grade ores. Higher-grade ores are usually put through more complex milling processes where higher recoveries justify the extra cost. The process chosen depends on the properties of the ore.
The final product is cathode copper.
Nickel Ores
The method is an acid heap leaching method like that of the copper method in that it utilises sulfuric acid instead of cyanide solution to dissolve the target minerals from crushed ore. The amount of sulfuric acid required is much higher than for copper ores (as high as 1,000 kg of acid per tonne of ore, but 500 kg is more common.) Nickel recovery from the leach solutions is much more complex than for copper and requires various stages of iron and magnesium removal, and the process produces both leached ore residue ("ripios") and chemical precipitates from the recovery plant (principally iron oxide residues, magnesium sulfate and calcium sulfate) in roughly equal proportions. Thus, a unique feature of nickel heap leaching is the need for a tailings disposal area.
The final product can be nickel hydroxide precipitates (NHP) or mixed metal hydroxide precipitates (MHP), which are then subject to conventional smelting to produce metallic nickel.
The method was originally patented by Australian miner BHP Billiton and is being commercialized by Cerro Matoso S.A. in Colombia (a wholly owned subsidiary of BHP Billiton), Vale in Brazil, and European Nickel PLC for the rock laterite deposits of Turkey, Talvivaara mine in Finland, Balkans, and the Philippines. There currently are no operating commercial scale nickel laterite heap leach operations, but there is a sulfide HL operating in Finland.
Uranium Ores
Heap leaching of uranium ores is similar to copper oxide heap leaching, also using dilute sulfuric acid. Rio Tinto is commercializing this technology in Namibia and Australia, the French nuclear power company Areva in Niger (two mines) and Namibia, and several other companies are studying its feasibility.
The final product is yellowcake and requires significant further processing to produce fuel-grade feed.
Heap Leaching Apparatus
While most mining companies have shifted from a previously accepted sprinkler method to the percolation of slowly dripping choice chemicals (cyanide or sulfuric acid) closer to the actual ore bed, heap leach pads have not changed too much throughout the years. There are still four main categories of pads: conventional, dump leach, Valley Fills, and on/off pads. Typically, each pad only has a single, geomembrane liner for each pad, with a minimum thickness of 1.5mm (usually it is thicker).
The simplest in design, conventional pads are used for mostly flat or gentle areas and hold thinner layers of crushed ore. Dump leach pads hold more ore and can usually handle a less flat terrain. Valley Fills are pads situated at valley bottoms or levels that can hold everything falling into it. On/off pads involve the use of putting significantly larger loads on the pads, and removing it and reloading it after every cycle.
Many of these mines, which previously had digging depths of about 15 meters, are digging deeper than ever before to mine materials (approximately 50 meters, sometimes more), which means that, in order to accommodate all of the ground being displaced, pads will have to hold higher weights from more crushed ore being contained in a smaller area. With that increase in buildup comes in potential for decrease in yield or ore quality, as well as potential either weak spots in the lining or areas of increased pressure buildup. This build up still has the potential to lead to punctures in the liner. As of 2004 cushion fabrics, which could reduce potential punctures and their leaking, were still being debated due to their tendency to increase risks if too much weight on too large a surface was placed on the cushioning. In addition, some liners, depending on their composition, may react with salts in the soil as well as acid from the chemical leaching to affect the successfulness of the liner. This can be amplified over time.
Heap Leaching Environmental Concerns
Heap leach mining works well for high concentrations of less ores, as less Earth needs to be ground onto leach pads in order to extract the same amount of materials. While yield is usually approximately 60-70%, there are significant amounts of damage to the surface environment. Yet the amount of overall harm caused by heap leaching is often lower than more traditional techniques, reducing costs to the process. It also requires less energy consumption to use these methods, which many consider to be an environmental alternative.
In some cases, waste materials from this process are transported to a facility to be treated. However, after a month (or more) of setting in a mat, there is often still more wait time involved in recovering the chemicals—both pregnant and excess—thus allowing for additional chemicals to potentially leach out of the pad into the soil below. This could possibly cause damage to the environment, which has a chance at contaminating surrounding bodies of water.
All in all, it is only slightly more environmentally friendly than in-situ mining, which involves leaving chemicals directly into the soil before pulling pregnant chemicals up. Still, most soil in heap leaching is seldom replaced or additionally treated. Evidence has also been found to show that there increased levels of erosion in mining sites with open chemicals, including these heap leach mines, which could be exacerbated through natural phenomena like storms and wind, or more serious occurrences like earthquakes.
The threat to ecosystem composition and biodiversity has been commented on repeatedly in terms of these mines, and it is noted that, while they do have a higher yield, they also have tendencies to accumulate wear and tear from being outside, and pose a threat to the immediate environment by crushing and dumping dirt that would otherwise have been left untouched. As noted earlier, with the reduction of readily available rare earth minerals, there has been increased in the amount of ore piled onto these pads, suggesting that there may come a time when the amount of ore dumped is not worth the amount of returned mineral collected. Therefore, alternatives need to be considered in the near future. Currently, depths are being mined faster than research can provide information about the effects of more ore on the system.
There is also very little study for long-term viability of liners, as this type of mining and the increased ore depths are still a relatively new field, especially given the changes in depths that have been put into practice. With the increase in weight, pressure, and chemicals put on this method of mining, as well as the already small level of knowledge regarding long term benefits, it is difficult to predict the extent of damage from previous leaks, as well as the durability of present day pads and mining sites.
Heap Leaching Legal Claims
Mining laws have been lobbied against in the United States for the past few decades, and many are only beginning to hear the environmentalist argument. The US General Mining Law of 1872 previous gave liberal rights to miners in terms of establishing and exploring claims, yet did not require any sort of environmental rehabilitation aspect to its process. Today, this is being disputed due to the number of environmental problems being found in heap leach sites, as well as the increase in scientific knowledge that could make mining more efficient and less costly as well. There has been much debate going into the levels of revision of the U.S. General Mining Law of 1872, including whether rehabilitation measures should be added to a decrease in the amount of liberal rights given to mining companies. Australia has already addressed much of this with the increased amounts of impact and externality knowledge required in any mining proposal endeavor.
However, as seen with many case studies, a simple way around these measures is the privatization of the land to be mined. In this case, environmental standards need to be made priority, as their effects spread beyond simple legal boundaries and into the ecosystems present at each mine, many of which are affecting poorer people less likely to speak up for their health and their environment.
Heap Leaching Cultural and Social Concerns
With the rise of the environmentalist movement has also come an increased appreciation for social justice, and mining has showed similar trends lately. Societies located near potential mining sites are at increased risk to be subjected to injustices as their environment is affected by the changes made to mined lands—either public or private—that could eventually lead to problems in social structure, identity, and physical health. Many have argued that by cycling mine power through local citizens, this disagreement can be alleviated, since both interest groups would have shared and equal voice and understanding in future goals. However, it is often difficult to match corporate mining interests with local social interests, and money is often a deciding factor in the successes of any disagreements. If communities are able to feel like they have a valid understanding and power in issues concerning their local environment and society, they are more likely to tolerate, and indeed encourage, the positive benefits that come with mining, as well as more effectively promote alternative methods to heap leach mining using their intimate knowledge of the local geography. Through increased dialogue and environmental legislation, many corporations and citizens hope to bridge the gap between interests in order to obtain the rare natural resources that most people depend on in daily life.
Dump leaching
Dump leaching is an industrial process to extract precious metals and copper from ores.
Dump leaching is similar to heap leaching, however in the case of dump leaching ore is taken directly from the mine and stacked on the leach pad without crushing where, in the case of gold and silver, the dump is irrigated with a dilute cyanide solution that percolates through the ore to dissolve gold and silver. The solution containing gold and silver exits the base of the dump, is collected and precious metals extracted. The resultant barren solution is recharged with additional cyanide and returned to the dump.
This method of leaching is usually suitable for low grade ores because it is very low cost. However, it operates with slow kinetics and may take up about 1 to 2 years to extract 50% of the desired mineral.
Tank and Vat leaching
In metallurgical processes tank leaching is a hydrometallurgical method of extracting valuable material (usually metals) from ore.
Tank vs. vat leaching
Tank leaching is usually differentiated from vat leaching on the following factors:
1. In tank leaching the material is ground sufficiently fine to form a slurry or pulp, which can flow under gravity or when pumped. In vat leaching typically a coarser material is placed in the vat for leaching, this reduces the cost of size reduction;
2. Tanks are typically equipped with agitators, baffles, gas introduction equipment designed to maintain the solids in suspension in the slurry, and achieve leaching. Vats usually do not contain “internal” equipment;
3. Tank leaching is typically continuous, while vat leaching is operated in a batch fashion, this is not always the case, and commercial processes using continuous vat leaching have been tested;
4. Typically the retention time required for vat leaching is more than that for tank leaching to achieve the same percentage of recovery of the valuable material being leached;
In a tank leach the slurry is moved, while in a vat leach the solids remain in the vat, and solution is moved.
Tank and vat leaching involves placing ore, usually after size reduction and classification, into large tanks or vats at ambient operating conditions containing a leaching solution and allowing the valuable material to leach from the ore into solution.
In tank leaching the ground, classified solids are already mixed with water to form a slurry or pulp, and this is pumped into the tanks. Leaching reagents are added to the tanks to achieve the leaching reaction. In a continuous system the slurry will then either overflow from one tank to the next, or be pumped to the next tank. Ultimately the “pregnant” solution is separated from the slurry using some form of liquid/solid separation process, and the solution passes on to the next phase of recovery.
In vat leaching the solids are loaded into the vat, once full the vat is flooded with a leaching solution. The solution drains from the tank, and is either recycled back into the vat or is pumped to the next step of the recovery process.
As mentioned previously tanks are equipped with agitators to keep the solids in suspension in the vats and improve the solid to liquid to gas contact. Agitation is further assisted by the use of tank baffles to increase the efficiency of agitation and prevent centrifuging of slurries in circular tanks.
Aside from chemical requirements several key factors influence extraction efficiency:
 Retention time - refers to the time spent in the leaching system by the solids. This is calculated as the total volumetric capacity of the leach tank/s divided by the volumetric throughput of the solid/liquid slurry. Retention time is commonly measured in hours for precious metals recovery. A sequence of leach tanks is referred to as a leach "train", and retention time is measured considering the total volume of the leach train. The desired retention time is determined during the testing phase, and the system is then designed to achieve this.
 Size - The ore must be ground to a size that exposes the desired mineral to the leaching agent (referred to as “liberation”), and in tank leaching this must be a size that can be suspended by the agitator. In vat leaching this is the size that is the most economically viable, where the recovery achieved as ore is ground finer is balanced against the increased cost of processing the material.
 Slurry density - The slurry density (percent solids) determines retention time. The settling rate and viscosity of the slurry are functions of the slurry density. The viscosity, in turn, controls the gas mass transfer and the leaching rate.
 Numbers of tanks - Agitated tank leach circuits are typically designed with no less than four tanks and preferably more to prevent short-circuiting of the slurry through the tanks.
 Dissolved gas - Gas is often injected below the agitator or into the vat to obtain the desired dissolved gas levels – typically oxygen, in some base metal plants sulphur dioxide may be required.
 Reagents - Adding and maintaining the appropriate amount of reagents throughout the leach circuit is critical to a successful operation. Adding insufficient quantities of reagents reduces the metal recovery but adding excess reagents increases the operating costs without recovering enough additional metal to cover the cost of the reagents.
The tank leaching method is commonly used to extract gold and silver from ore.
In-situ leach
In-situ leaching (ISL), also called in-situ recovery (ISR) or solution mining, is a leaching process used to recover minerals such as copper and uranium through boreholes drilled into a deposit, in situ.
The process initially involves drilling of holes into the ore deposit. Explosive or hydraulic fracturing may be used to create open pathways in the deposit for solution to penetrate. Leaching solution is pumped into the deposit where it makes contact with the ore. The solution bearing the dissolved ore content is then pumped to the surface and processed. This process allows the extraction of metals and salts from an ore body without the need for conventional mining involving drill-and-blast, open-cut or underground mining.
In-situ leach mining involves pumping of a leachate solution into the ore body via a borehole, which circulates through the porous rock dissolving the ore and is extracted via a second borehole.
The leachate solution varies according to the ore deposit: for salt deposits the leachate can be fresh water into which salts can readily dissolve. For copper, acids are generally needed to enhance solubility of the ore minerals within the solution. For uranium ores, the leachate may be acid or sodium bicarbonate.
Soluble salts
In-situ leaching is widely used to extract deposits of water-soluble salts such as sylvite (potash), halite (rock salt, sodium chloride), and sodium sulfate. It has been used in the US state of Colorado to extract nahcolite (sodium bicarbonate). In-situ leaching is often used when the deposits are too deep, or the beds too thin for conventional underground mining.
Uranium
Solutions used to dissolve uranium ore are either acid (sulfuric acid or less commonly nitric acid) or carbonate (sodium bicarbonate, ammonium carbonate, or dissolved carbon dioxide). Dissolved oxygen is sometimes added to the water to mobilize the uranium. ISL of uranium ores started in the United States and the Soviet Union in the early 1960s. The first uranium ISL in the US was in the Shirley Basin in the state of Wyoming, which operated from 1961-1970 using sulfuric acid. Since 1970, all commercial-scale ISL mines in the US have used carbonate solutions.
In-situ recovery involves the extraction of uranium-bearing water (grading as low as 0.05% U3O8). The extracted uranium solution is then filtered through resin beads. Through an ion exchange process, the resin beads attract uranium from the solution. Uranium loaded resins are then transported to a processing plant, where U3O8 is separated from the resin beads and yellowcake is produced. The resin beads can then be returned to the ion exchange facility where they are reused.
At the end of 2008 there were four in-situ leaching uranium mines operating in the United States, all using sodium bicarbonate. ISL produces 90% of the uranium mined in the US.
Significant ISL mines are operating in Kazakhstan and Australia. ISL mining accounted for 41% of the world's uranium production in 2010.
Copper
In-situ leaching of copper was done by the Chinese by 977 AD, and perhaps as early as 177 BC. Copper is usually leached using acid (sulfuric acid or hydrochloric acid), then recovered from solution by solvent extraction electrowinning (SX-EW) or by chemical precipitation.
Ores most amenable to leaching include the copper carbonates malachite and azurite, the oxide tenorite, and the silicate chrysocolla. Other copper minerals, such as the oxide cuprite and the sulfide chalcocite may require addition of oxidizing agents such as ferric sulfate and oxygen to the leachate before the minerals are dissolved. The ores with the highest sulfide contents, such as bornite and chalcopyrite will require more oxidants and will dissolve more slowly. Sometimes oxidation is speeded by the bacteria Thiobacillus ferrooxidans, which feeds on sulfide compounds.
Copper ISL is often done by stope leaching, in which broken low-grade ore is leached in a current or former conventional underground mine. The leaching may take place in backfilled stopes or caved areas. In 1994, stope leaching of copper was reported at 16 mines in the US. At the San Manuel mine in the US state of Arizona, ISL, underground mining, and open-pit mining were being done simultaneously in different parts of the same ore body.
Gold
In-situ leaching has not been used on a commercial scale for gold mining. A three-year pilot program was undertaken in the 1970s to in-situ leach gold ore at the Ajax mine in the Cripple Creek district in the US, using a chloride and iodide solution. After obtaining poor results, perhaps because of the complex telluride ore, the test was halted.
Environmental Concerns: ISL Mining of Uranium
In the USA legislation requires that the water quality in the affected aquifer be restored so as to enable its pre-mining use. Usually this is potable water or stock water (usually less than 500 ppm total dissolved solids), and while not all chemical characteristics can be returned to those pre-mining, the water must be usable for the same purposes as before. Often it needs to be treated by reverse osmosis, giving rise to a problem in disposing of the concentrated brine stream from this. The usual radiation safeguards are applied at an ISL Uranium mining operation, despite the fact that most of the ore body’s radioactivity remains well underground and there is hence minimal increase in radon release and no ore dust. Employees are monitored for alpha radiation contamination and personal dosimeters are worn to measure exposure to gamma radiation. Routine monitoring of air, dust and surface contamination are undertaken.
The leaching liquid used for in-situ leaching contains the leaching agent ammonium carbonate for example, or - particularly in Europe - sulfuric acid. This method can only be applied if the uranium deposit is located in porous rock, confined in impermeable rock layers.
The advantages of this technology are:
 Reduced hazards for the employees from accidents, dust, and radiation,
 Low cost, no need for large uranium mill tailings deposits.
The disadvantages of the in-situ leaching technology are:
 Risk of spreading of leaching liquid outside of the uranium deposit, involving subsequent groundwater contamination
 Unpredictable impact of the leaching liquid on the rock of the deposit, and
 the impossibility of restoring natural groundwater conditions after completion of the leaching operations.
Moreover, in-situ leaching releases considerable amounts of radon, and produces certain amounts of waste slurries and waste water during recovery of the uranium from the liquid.
After termination of an in-situ leaching operation, the waste slurries produced must be safely disposed, and the aquifer, contaminated from the leaching activities, must be restored. Groundwater restoration is a very tedious process that is not yet fully understood. So far, it is not possible to restore groundwater quality to previous conditions.
Newbery-Vautin chlorination process
The Newbery-Vautin chlorination process is a process to extract gold from its ore using chlorination developed by James Cosmo Newbery and Claude Vautin.
The process of extracting gold from ores by absorption of the precious metal in chlorine gas, from which it is reduced to a metallic state, is not a very new discovery. It was first introduced by Karl Friedrich Plattner around 1848, and at that time promised to revolutionize the processes for gold extraction. By degrees it was found that only a very clever chemist could work this process with practically perfect results, for many reasons. Lime and magnesia might be contained in the quartz, and would be attacked by the chlorine. These consume the reagents without producing any results, earthy particles would settle and surround the small gold and prevent chlorination, then lead and zinc or other metals in combination with the gold would also be absorbed by the chlorine; or, again, from some locally chemical peculiarity in the water or the ore, gold held in solution by the water might be again precipitated in the tailings before filtration was complete, and thus be lost. Henderson, Clark, De Lacy, Mears, and Deacon, all introduced improvements, or what were claimed to be improvements, on Plattner, but these chiefly failed because they did not cover every particular variety of case which gold extraction presented. Therefore, where delicate chemical operations were necessary for success, practice generally failed from want of knowledge on the part of the operator, and many times extensive plants have been pronounced useless from this cause alone. Hence it is not to be wondered that processes requiring such care and uncommon knowledge are not greatly in favor.
Claude Theodore James Vautin, a gentleman possessed of much practical experience of gold mining and extraction in Queensland, Australia, together with James Cosmo Newbery, analytical chemist to the government of Victoria, have developed a process which they claim to combine all the advantages of the foregoing methods, and by the addition of certain improvements in the machinery and mode of treatment to overcome the difficulties which have hitherto prevented the general adoption of the chlorination process.
The materials for treatment—crushed and roasted ore, or tailings, as the case may be—are put into the hopper above the revolving barrel, or chlorinator. This latter is made of iron, lined with wood and lead, and sufficiently strong to bear a pressure of 100 lb. to the square inch, its capacity being about 30 cwt. of ore. The charge falls from the hopper into the chlorinator. Water and chlorine-producing chemicals are added—generally sulphuric acid and chloride of lime—the manhole cover is replaced and screwed down so as to be gas tight. On the opposite side of the barrel there is a valve connected with an air pump, through which air to about the pressure of four atmospheres is pumped in, to liquefy the chlorine gas that is generated, after which the valve is screwed down. The barrel is then set revolving at about ten revolutions a minute, the power being transmitted by a friction wheel. According to the nature of the ore, or
the size of the grains of gold, this movement is continued from one to four hours, during which time the gold, from combination with the chlorine gas, has formed a soluble gold chloride, which has all been taken up by the water in the barrel. The chlorinator is then stopped, and the gas and compressed air allowed to escape from the valve through a rubber hose into a vat of lime water. This is to prevent the inhalation of any chlorine gas by the workmen. The manhole cover is now removed and the barrel again set revolving, by which means the contents are thrown automatically into the filter below. This filter is an iron vat lined with lead. It has a false bottom, to which is connected a pipe from a vacuum pump working intermittently. As soon as all the ore has fallen from the chlorinator into the filter, the pump is set going, a partial vacuum is produced in the chamber below the false bottom in the filter, and very rapid filtration results. By this means all the gold chlorides contained in the wet ore may be washed out, a continual stream being passed through it while filtration is going on. The solution running from the filter is continually tested, and when found free from gold, the stream of water is stopped, as is also the vacuum pump. The filter is then tipped up into a truck below, and the tailings run out to the waste heap. The process of washing and filtration occupies about an hour, during which time another charge may be in process of treatment in the chlorinator above. The discharge from the filter and the washings are run into a vat, and from this they are allowed to pass slowly through a tap into a charcoal filter. During the passage of the liquid through the charcoal filter, the chloride of gold is decomposed and the gold is deposited on the charcoal, which, when fully charged, is burnt, the ashes are fused with borax in a crucible, and the gold is obtained.
Gold cyanidation
Gold cyanidation (also known as the cyanide process or the MacArthur-Forrest process) is a metallurgical technique for extracting gold from low-grade ore by converting the gold to a water soluble coordination complex. It is the most commonly used process for gold extraction. Production of reagents for mineral processing to recover gold, copper, zinc and silver represents approximately 13% of cyanide consumption globally, with the remaining 87% of cyanide used in other industrial processes such as plastics, adhesives, and pesticides. Due to the highly poisonous nature of cyanide, the process is controversial and its usage is banned in a number of countries and territories.
In 1783 Carl Wilhelm Scheele discovered that gold dissolved in aqueous solutions of cyanide. He had earlier discovered cyanide salts. Cyanide was not applied to extraction of gold ores until 1887, when the MacArthur-Forrest Process was developed in Glasgow, Scotland by John Stewart MacArthur, funded by the brothers Dr Robert and Dr William Forrest.
The chemical reaction for the dissolution of gold, the "Elsner Equation", follows:
4 Au + 8 NaCN + O2 + 2 H2O → 4 Na[Au(CN)2] + 4 NaOH
In this redox process, oxygen removes, via a two-step reaction, one electron from each gold atom to form the complex Au(CN)2- ion.
The ore is comminuted using grinding machinery. Depending on the ore, it is sometimes further concentrated by froth flotation or by centrifugal (gravity) concentration. Water is added to produce a slurry or pulp. The alkaline ore slurry can be combined with a solution of sodium cyanide or potassium cyanide, however many operations utilize calcium cyanide, which is more cost effective.
To prevent the creation of toxic hydrogen cyanide during processing, lime (calcium hydroxide) or soda (sodium hydroxide) is added to the extracting solution to ensure that the acidity during cyanidation is maintained over pH 10.5 - strongly alkaline. Lead nitrate can improve gold leaching speed and quantity recovered, particularly in processing partially oxidized ores.
Oxygen is one of the reagents consumed during cyanidation, and a deficiency in dissolved oxygen slows leaching rate. Air or pure oxygen gas can be purged through the pulp to maximize the dissolved oxygen concentration. Intimate oxygen-pulp contactors are used to increase the partial pressure of the oxygen in contact with the solution, thus raising the dissolved oxygen concentration much higher than the saturation level at atmospheric pressure. Oxygen can also be added by dosing the pulp with hydrogen peroxide solution.
In some ores, particularly those that are partially sulfidized, aeration (prior to the introduction of cyanide) of the ore in water at high pH can render elements such as iron and sulfur less reactive to cyanide, and therefore the gold cyanidation process more efficient. Specifically, the oxidation of iron to iron (III) oxide and subsequent precipitation as iron hydroxide minimizes loss of cyanide from the formation of ferrous cyanide complexes. The oxidation of sulfur compounds to sulfate ions avoids the consumption of cyanide to thiocyanate (SCN-) byproduct.
In order of decreasing economic efficiency, the common processes for recovery of the solubilized gold from solution are (certain processes may be precluded from use by technical factors):
 Carbon in pulp
 Electrowinning
 Merrill-Crowe process
Cyanide remediation processes
The various species of cyanide that remain in tails streams from gold plants are potentially toxic, and on some operations the waste streams are processed through a detoxification process prior to tails deposition. This reduces the concentrations of these cyanide compounds,
but does not completely eliminate them from the stream. The two major processes utilised are the INCO licenced process or the Caro’s acid process. Both processes utilise oxidants to oxidise cyanide to cyanate, which is not as toxic as the cyanide ion, and which can then react to form carbonates and ammonia. The Inco process can typically reduce cyanide concentrations to below 50 mg/L, while the Caro’s acid process can reduce cyanide levels to between 10 and 50 mg/L, with the lower concentrations achievable in solution streams rather than slurries. Hydrogen peroxide and alkaline chlorination can also be used, although these are typically less common.
One of the alternative oxidants for the degradation of cyanides that has been attracting industrial interest is Caro’s acid – peroxomonosulphuric acid (H2SO5). Caro’s acid converts cyanide to cyanate. Cyanate then hydrolyses in the water to ammonium and carbonate ions. The Caro's acid process is able to achieve discharge levels of WAD below 50 mg/L, which is generally suitable for discharge to tailings. Generally, the best application of this process is with tailings slurries containing low to moderate initial levels of cyanide and when treated cyanide levels of less than about 10 to 50 mg/L are required.
Over 90 mines worldwide now use an Inco SO2/air detoxification circuit to convert cyanide to the much less toxic cyanate before waste is discharged to a tailings pond. Typically, this process blows compressed air through the tailings while adding sodium metabisulfite which releases SO2, lime to maintain the pH at around 8.5, and copper sulfate as a catalyst if there is insufficient copper in the ore extract. This procedure can reduce concentrations of "Weak Acid Dissociable" (WAD) cyanide to below the 10 ppm mandated by the EU's Mining Waste Directive. This level compares to the 66-81 ppm free cyanide and 500-1000 ppm total cyanide in the pond at Baia Mare. Remaining free cyanide degrades in the pond, while cyanate ions hydrolyse to ammonium. Recent studies show that residual cyanide trapped in the gold-mine tailings causes persistent release of toxic metals (e.g. mercury) into the groundwater and surface water systems.
Effects on the environment
Despite being used in 90% of gold production, gold cyanidation is controversial due to the toxic nature of cyanide. Although aqueous solutions of cyanide degrade rapidly in sunlight, the less-toxic products, such as cyanates and thiocyanates, may persist for some years. The famous disasters have killed few people — humans can be warned not to drink or go near polluted water — but cyanide spills can have a devastating effect on rivers, sometimes killing everything for several miles downstream. However, the cyanide is soon washed out of river systems and, as long as organisms can migrate from unpolluted areas upstream, affected areas can soon be repopulated. In the Someș River below Baia Mare, the plankton returned to 60% of normal within 16 days of the spill. Such spills have prompted fierce protests at new mines that involve
use of cyanide, such as Roşia Montană in Romania, Lake Cowal in Australia, Pascua Lama in Chile, and Bukit Koman in Malaysia.
2. Amalgamation
An amalgam is a substance formed by the reaction of mercury with another metal. Almost all metals can form amalgams with mercury, a notable exception being iron. Silver-mercury amalgams are important in dentistry, and gold-mercury amalgam is used in the extraction of gold from ore.
Mercury has been used in the gold and silver mining methods because of the convenience and ease with which mercury will amalgamate with them. In gold placer mining, in which minute specks of gold are washed from sand or gravel deposits, mercury was often used to separate the gold from other heavy minerals.
After all of the practical metal had been taken out from the ore, the mercury was dispensed down a long copper trough, which formed a thin coating of mercury on the exterior. The waste ore was then transferred down the trough, and any gold in the waste was amalgamated with the mercury. This coating would sometimes get scraped off and refined to get rid of the mercury, leaving behind somewhat high purity gold.
Mercury amalgamation was first useful to silver ores with the development of the patio process in Mexico in 1557. There were also additional amalgamation processes that were created for processing silver ores, including pan amalgamation and the Washoe process.
Gold extraction (mining)
This has proved effective where gold fines ("flour gold") would not be extractable from ore using hydro-mechanical methods. Large amounts of mercury were used in placer mining, where deposits composed largely of decomposed granite slurry were separated in long runs of "riffle boxes", with mercury dumped in at the head of the run. The amalgam formed is a heavy solid mass of dull gray color. (The use of mercury in 19th century placer mining in California, now prohibited, has caused extensive pollution problems in riverine and estuarine environments, ongoing to this day.) Sometimes substantial slugs of amalgam are found in downstream river and creek bottoms by amateur wet-suited miners seeking gold nuggets with the aid of an engine-powered water vacuum mounted on a float.
Where stamp mills were used to crush gold-bearing ore to fines, a part of the extraction process involved the use of mercury-wetted copper plates, over which the crushed fines were washed. A periodic scraping and re-mercurizing of the plate resulted in amalgam for further processing.
Amalgam obtained by either process was then heated in a distillation retort, recovering the mercury for reuse and leaving behind the gold. As this released mercury vapors to the atmosphere, the process could induce adverse health effects and long term pollution.
Today, mercury amalgamation has been replaced by other methods to recuperate gold and silver from ore in developed nations. Hazards of mercurial toxic waste have played a major role in the phasing out of the mercury amalgamation processes. However mercury amalgamation is still regularly used by small-scale gold placer miners (often illegally) & particularly in developing countries.
Mercury salts are, compared to mercury metal and amalgam, highly toxic due to their solubility in water. The presence of these salts in water can be detected with a probe that uses the readiness of mercury ions to form an amalgam with copper. A nitric acid solution of salts under investigation is applied to a piece of copper foil and any mercury ions present will leave spots of silvery-coloured amalgam. Silver ions leave similar spots but are easily washed away, making this a means of distinguishing silver from mercury.
The redox reaction involved where mercury oxidizes the copper is:
Hg2+ + Cu → Hg + Cu2+
Patio process
The patio process was a process used to extract silver from ore. The process was invented by Bartolomé de Medina in Pachuca, New Spain (Mexico), in 1554. The patio process was the first process to use mercury amalgamation to recover silver from ore. It replaced smelting as the primary method of extracting silver from ore at Spanish colonies in the Americas. Other amalgamation processes were later developed, importantly the pan amalgamation process, and its variant, the Washoe process. The silver separation process generally differed from gold parting and gold extraction, although amalgamation with mercury was also sometimes used to extract gold.
Basic elements of the patio process
Silver ores were crushed (typically either in "arrastras" or stamp mills) to a fine slime which was mixed with salt, water, magistral (essentially an impure form of copper sulfate), and mercury, and spread in a 1-to-2-foot-thick (0.30 to 0.61 m) layer in a patio, (a shallow-walled, open enclosure). Horses were driven around on the patio to further mix the ingredients. After weeks of mixing and soaking in the sun, a complex reaction converted the silver to native metal, which formed an amalgam with the mercury and was recovered. The amount of salt and copper sulfate varied from one-quarter to ten pounds of one or the other, or both, per ton of ore treated. The decision of how much of each ingredient to add, how much mixing was needed,
and when to halt the process depended on the skill of an azoguero (English: quicksilver man). The loss of mercury in amalgamation processes is generally one to two times the weight of silver recovered.
The patio process was the first form of amalgamation. However, it is unclear whether this process or a similar process—in which amalgamation occurred in heated vats rather than open patios—was the predominant form of amalgamation in New Spain, as the earliest known illustration of the patio process dates from 1761. There is substantial evidence that both processes were used from an early date in New Spain, while open patios were never adopted in Peru. Both processes required that ore be crushed and refiners quickly established mills to process ore once amalgamation was introduced. By the seventeenth century, water-powered mills became dominant in both New Spain and Peru.
Due to amalgamation's reliance upon mercury, an expansion of mercury production was central to the expansion of silver production. From shortly after the invention of mercury amalgamation to the end of the colonial period, the Spanish crown maintained a monopoly on mercury production and distribution, ensuring a steady supply of royal income. Fluctuations in mercury prices generally resulted in corresponding increases and decreases in silver production.
Pan amalgamation
The Pan amalgamation process is a method to extract silver from ore, using mercury. The process was widely used from 1609 through the 19th century; it is no longer used.
One drawback of the patio process was the long treatment time, usually weeks. Alvaro Alonso Barba invented the faster pan process (in Spanish the cazo or fondo process) in 1609 in Potosí, Bolivia, in which ore was mixed with salt and mercury (and sometimes copper(II) sulfate) and heated in shallow copper vessels. The treatment time was reduced to 10 to 20 hours. Whether patio or pan amalgamation was used at a particular location often depended on climate (warmer conditions speeded the patio process) and the availability and cost of fuel to heat the pans.
The amount of salt and copper(II) sulfate varied from one-quarter to ten pounds of one or the other, or both, per ton of ore treated. The loss of mercury in amalgamation processes was generally one to two times the weight of silver recovered.
Washoe process
The Washoe process, a variation of pan amalgamation, was developed in the 1860s by Almarin Paul and others, to work the ore from the Comstock Lode in Nevada, United States (Washoe
was an early name for the area). In the Washoe process, the copper pans were replaced by iron tanks with mechanical agitators. Each tank ("pan") was circular, and commonly held 1,200 to 1,500 pounds of ore that had been crushed to sand size. Water was added to make a pulp, and 60 to 70 pounds of mercury, along with one-half to three pounds each of salt (sodium chloride) and bluestone (copper(II) sulfate) were also added. A circular iron plate called a muller was mounted on a vertical shaft and lowered into the tank, and was rotated to provide both agitation and additional grinding. Heat was delivered to the tanks by steam pipes. The iron filings worn from the muller and pan proved to be an essential ingredient in the process.
Reese River process
A variation of the Washoe process was developed in the Reese River mining district around Austin, Nevada. The Washoe process was found not to work well for ores with arsenic or antimony sulfides, or with galena or sphalerite. In 1869, Carl A. Stetefeldt of Reno found that roasting the ore with salt converted the silver sulfides to silver chlorides, which could then be recovered in amalgamation pans. The process was introduced in the Reese River District in 1879, with great success.
ELECTROMETALLURGICAL ORE REDUCTION METHODS
Electrometallurgy is the field concerned with the processes of metal electrodeposition. There are four categories of these processes, but our focus here will be on electrowinning, the extraction of metal from ores.
1. Electrowinning
Electrowinning, also called electroextraction, is the electrodeposition of metals from their ores that have been put in solution or liquefied. Electrorefining uses a similar process to remove impurities from a metal. Both processes use electroplating on a large scale and are important techniques for the economical and straightforward purification of non-ferrous metals. The resulting metals are said to be electrowon.
In electrowinning, a current is passed from an inert anode through a liquid leach solution containing the metal so that the metal is extracted as it is deposited in an electroplating process onto the cathode. In electrorefining, the anodes consist of unrefined impure metal, and as the current passes through the acidic electrolyte the anodes are corroded into the solution so that the electroplating process deposits refined pure metal onto the cathodes.
Applications
The most common electrowon metals are lead, copper, gold, silver, zinc, aluminium, chromium, cobalt, manganese, and the rare-earth and alkali metals. For aluminium, this is the only production process employed. Several industrially important active metals (which react strongly with water) are produced commercially by electrolysis of their pyrochemical molten salts. Experiments using electrorefining to process spent nuclear fuel have been carried out. Electrorefining may be able to separate heavy metals such as plutonium, caesium, and strontium from the less-toxic bulk of uranium. Many electroextraction systems are also available to remove toxic (and sometimes valuable) metals from industrial waste streams.
Process
Most metals occur in nature in their oxidized form (ores) and thus must be reduced to their metallic forms. The ore is dissolved following some preprocessing in an aqueous electrolyte or in a molten salt and the resulting solution is electrolyzed. The metal is deposited on the cathode (either in solid or in liquid form), while the anodic reaction is usually oxygen evolution. Several metals are naturally present as metal sulfides; these include copper, lead, molybdenum, cadmium, nickel, silver, cobalt, and zinc. In addition, gold and platinum group metals are associated with sulfidic base metal ores. Most metal sulfides or their salts are electrically conductive and this allows electrochemical redox reactions to efficiently occur in the molten state or in aqueous solutions.
Fig. 3: Apparatus for electrolytic refining of copper
Some metals, including arsenic and nickel do not electrolyze out but remain in the electrolyte solution. These are then reduced by chemical reactions to refine the metal. Other metals, which during the processing of the target metal have been reduced but not deposited at the cathode, sink to the bottom of the electrolytic cell, where they form a substance referred to as anode sludge or anode slime. The metals in this sludge can be removed by standard pyrorefining methods.
Because metal deposition rates are related to available surface area, maintaining properly working cathodes is important. Two cathode types exist, flat-plate and reticulated cathodes, each with its own advantages. Flat-plate cathodes can be cleaned and reused, and plated metals recovered. Reticulated cathodes have a much higher deposition rate compared to flat-plate cathodes. However, they are not reusable and must be sent off for recycling. Alternatively, starter cathodes of pre-refined metal can be used, which become an integral part of the finished metal ready for rolling or further processing.
ORE REFINING OR PURIFICATION METHODS
The metal obtained after reduction process still contains some impurities which can be removed by applying the following methods.
PYROMETALLURGICAL METAL PURIFICATION METHODS
1. Cupellation
Cupellation is a refining process in metallurgy, where ores or alloyed metals are treated under high temperatures and controlled operations to separate noble metals, like gold and silver, from base metals like lead, copper, zinc, arsenic, antimony or bismuth, present in the ore. The process is based on the principle that precious metals do not oxidise or react chemically, unlike the base metals; so when they are heated at high temperatures, the precious metals remain apart and the others react forming slags or other compounds.
Since the Early Bronze Age, the process was used to obtain silver from smelted lead ores. By the Middle Ages and the Renaissance, cupellation was one of the most common processes for refining precious metals. By then, fire assays were used for assaying minerals, that is, testing fresh metals such as lead and recycled metals to know their purity for jewelry and coin making. Cupellation is still in use today.
Process
Large scale cupellation
Native silver is a rare element, although it exists as such. It is usually found in nature combined with other metals, or in minerals that contain silver compounds, generally in the form of sulfides such as galena (lead sulfide) or cerussite (lead carbonate). So the primary production of silver requires the smelting and then cupellation of argentiferous lead ores.
Lead melts at 327°C while silver melts at 960°C; when mixed, as in galena, the most common argentiferous lead ore, they have to be smelted at high temperatures in a reducing atmosphere
to produce argentiferous lead. The alloy is melted again at the high temperature of 900°C to 1000°C in a hearth or blast furnace, where air flow makes possible the oxidation of the lead. The lead oxidises to lead oxide (PbO) known as litharge, which captures the oxygen from the other metals present, while silver and gold remain on top of the liquid litharge. The latter is removed or absorbed by capillary action into the hearth linings. This chemical reaction may be viewed as:
(Ag+Cu) + Pb + O2 → (CuO+PbO) + Ag
The base of the hearth was dug in the form of a saucepan, and covered with an inert and porous material rich in calcium or magnesium such as shells, lime, or the ash from burning wood or bones. The lining had to be calcareous because lead reacts with silica (clay compounds) to form viscous lead silicate that prevents the needed absorption of litharge, whereas calcareous materials do not react with lead. Some of the litharge evaporates, and the rest is absorbed by the porous earth lining to form "litharge cakes".
Litharge cakes are usually circular or concavo-convex, about 15 cm in diameter. They are the most common archaeological evidence of cupellation in the Early Bronze Age. By their chemical composition, archaeologists can tell what kind of ore was treated, its main components, and the chemical conditions used in the process. This permits insights about production process, trade, social needs or economic situations.
Small scale cupellation
Small scale cupellation is based on the same principle as the one done in a cupellation hearth; the main difference lies in the amount of material to be tested or obtained. The minerals have to be crushed, roasted and smelted to concentrate the metallic components in order to separate the noble metals. By the Renaissance the use of the cupellation processes was diverse: assay of ores from the mines, testing the amount of silver in jewels or coins or for experimental purposes. It was carried out in small shallow recipients known as cupels.
As the main purpose of small scale cupellation was to assay and test minerals and metals, the material to be tested had to be carefully weighed. The assays were made in the cupellation or assay furnace, which needed to have windows and bellows to ascertain that the air oxidises the lead, as well as to be sure and prepared to take away the cupel when the process was over. Pure lead had to be added to the material being tested to guarantee the further separation of the impurities. After the litharge had been absorbed by the cupel, buttons of silver were formed and settled in the middle of the cupel. If the alloy also contained a certain amount of gold, it settled with the silver and both had to be separated by parting.
Cupels
The primary tool for small scale cupellation was the cupel. Cupels were manufactured in a very careful way. They used to be small vessels shaped in the form of an inverted truncated cone, made out of bone ashes. According to Georg Agricola, the best material was obtained from burned horns of deer although fish spines could work well. Ashes have to be ground into a fine and homogeneous powder and mixed with some sticky substance to mould the cupels. Moulds were made out of brass with no bottoms so that the cupels could be taken off. A shallow depression in the centre of the cupel was made with a rounded pestle. Cupel sizes depend on the amount of material to be assayed. This same shape has been maintained until the present.
Archaeological investigations as well as archaeometallurgical analysis and written texts from the Renaissance have demonstrated the existence of different materials for their manufacture; they could be made also with mixtures of bones and wood ashes, of poor quality, or moulded with a mixture of this kind in the bottom with an upper layer of bone ashes. Different recipes depend on the expertise of the assayer or on the special purpose for which it was made (assays for minting, jewelry, testing purity of recycled material or coins). Archaeological evidence shows that at the beginnings of small scale cupellation, clay cupels were used.
2. Parkes process
The Parkes process is a pyrometallurgical industrial process for removing silver from lead, during the production of bullion. It is an example of liquid-liquid extraction.
The process takes advantage of two liquid-state properties of zinc. The first is that zinc is immiscible with lead, and the other is that silver is 3000 times more soluble in zinc than it is in lead. When zinc is added to liquid lead that contains silver as a contaminant, the silver preferentially migrates into the zinc. Because the zinc is immiscible in the lead it remains in a separate layer and is easily removed. The zinc-silver solution is then heated until the zinc vaporizes, leaving nearly pure silver. If gold is present in the liquid lead, it can also be removed and isolated by the same process.
3. Polling
Polling is a method employed in the purification of copper which contains cuprous oxide as an impurity. The impure metal is melted and green wooden poles are used to agitate the molten impure copper. The heat of the copper makes the pole emit a gas which reduces the cuprous oxide to copper.
4. Distillation method
Volatile metals (Hg, Zn, Cd) are easily purified by distillation. The impure metals are heated in a retort and vapours of volatile metals are collected and condensed in a receiver leaving behind non-volatile impurities in the retort.
5. Zone refining of fractional crystallization
This method is employed to get metals of very high purity (Ge, Si, B, Ga, In). The method is based on the difference in solubility of impurities of molten and solid state of the metal. A movable heater is allowed to move across the impure metals rod from one end to the other end. The pure metal crystallises while the impurities pass on to the adjacent melted zone.
6. Mond’s Process
This method is employed for purification of nickel. Impure nickel is converted into volatile nickel carbonyl by reaction of CO at 60-80oC. Nickel carbonyl decomposes at 180oC to form pure nickel and CO.
Ni + 4CO Ni(CO)4 Ni +4CO
HYDROMETALLURGICAL PURIFICATION METHODS
1. Leaching
Bayer process
The Bayer process is the principal industrial means of refining bauxite to produce alumina (aluminum oxide). Bauxite, the most important ore of aluminum, contains only 30–54% aluminum oxide, (alumina), Al2O3, the rest being a mixture of silica, various iron oxides, and titanium dioxide. The aluminum oxide must be purified before it can be refined to aluminum metal.
Process
In the Bayer process, bauxite is digested by washing with a hot solution of sodium hydroxide, NaOH, at 175°C. This converts the aluminum oxide in the ore to sodium aluminate, 2NaAl(OH)4, according to the chemical equation:
Al2O3 + 2NaOH + 3H2O → 2NaAl(OH)4
The other components of bauxite do not dissolve. The solution is clarified by filtering off the solid impurities. The mixture of solid impurities is called red mud, and presents a disposal problem. Next, the alkaline solution is cooled, and aluminum hydroxide precipitates as a white, fluffy solid:
NaAl(OH)4 → Al(OH)3 + NaOH
Then, when heated to 980°C (calcined), the aluminum hydroxide decomposes to aluminum oxide, giving off water vapor in the process:
2Al(OH)3 → Al2O3 + 3H2O
A large amount of the aluminum oxide so produced is then subsequently smelted in the Hall–Héroult process in order to produce aluminum.
ELECTOMETALLURGICAL PURIFICATION METHODS
1. Electrolysis
The impure metal is made the anode while a thin sheet of pure metal acts as the cathode. The electrolytic solution consists of an aqueous solution of salt or a complex of the metal. On passing the current, pure metal is deposited on the cathode and an equivalent amount of the
metal gets dissolved from anode. The soluble impurities pass into the solution and the insoluble impurities collect below the anode as anode mud.
The Hall–Héroult process
The Hall–Héroult process is the major industrial process for the production of aluminum. It involves dissolving alumina in molten cryolite, and electrolyzing the molten salt bath to obtain pure aluminum metal.
Fig. 5: A Hall–Héroult industrial cell
Process
Aluminum cannot be produced by the electrolysis of an aluminum salt dissolved in water because of its high reactivity: the water (especially, hydronium ions which are its natural constituent) readily oxidizes elemental aluminum. The reduction of Al3+ is done by electrolysis of a molten aluminum salt. This is a water-free medium.
Alumina, Al2O3, is dissolved in an industrial carbon-lined vat of molten cryolite, Na3AlF6 (sodium hexafluoroaluminate), called a "cell". Aluminum oxide has a melting point of over 2,000 °C while pure cryolite has a melting point of 1,012 °C. With a small percentage of alumina
dissolved in it, cryolite has a melting point of about 1,000 °C. Some aluminum fluoride, AlF3 is also added into the process to reduce the melting point of the cryolite-alumina mixture.
The molten mixture of cryolite, alumina, and aluminum fluoride is then electrolyzed by passing a direct electric current through it. The electrochemical reaction causes liquid aluminum metal to be deposited at the cathode as a precipitate, while the oxygen from the alumina combines with carbon from the anode to produce carbon dioxide, CO2. An electric potential of three to five volts is needed to drive the reaction, and the rate of production is proportional to the electric current. An industrial-scale smelter typically consumes hundreds of thousands of amperes for each cell.
The oxidation of the carbon anode reduces the required voltage across each cell, increasing the electrical efficiency, at a cost of continually replacing the carbon electrodes with new ones, and also the cost of releasing carbon dioxide into the atmosphere. Hundreds of Hall–Héroult cells are usually connected electrically in series, and they are supplied with direct current (DC) from a single set of rectifiers that convert alternating current (AC) supplied to the factory into direct current. The very high electric current is supplied to the cells through heavy, low electrical resistance metal busbars made of pure aluminum or copper. The cells are electrically heated to reach the operating temperature with this current, and the anode regulator system varies the current passing through the cell by raising or lowering the anodes and changing the cell's resistance. If needed any cell can be bypassed by shunt busbars.
The liquid aluminum is taken out with the help of a siphon operating with a vacuum, in order to avoid having to use extremely high temperature valves and pumps. The liquid aluminum then may be transferred in batches or via a continuous hot flow line to a location where it is cast into aluminum ingots. The aluminum can either be cast into the form of final cast-aluminum products, or the ingots can be sent elsewhere such as a rolling mill to be pressed into sheets, or a wire-drawing mill producing aluminum wires and cables.
While solid cryolite is denser than solid aluminum at room temperature, the liquid aluminum product is denser than the molten cryolite at temperatures around 1,000 °C, and the aluminum sinks to the bottom of the electrolytic cell, where it is periodically collected. The tops and sides of the cells are covered with layers of solid cryolite which also act as thermal insulation. The unavoidable electric resistance within each cell produces sufficient heat to keep the cryolite-alumina mixture molten.
With the percentage of aluminum dissolved in each cell being depleted by the electrolysis in the molten cryolite, additional alumina is continually dropped into the cells to maintain the required level of alumina. Whenever a solid crust forms across the surface of the molten
cryolite-alumina, this crust is broken from time to time to allow the added alumina to fall into the molten cryolite and dissolve there.
The electrolysis process produces exhaust which escapes into the fume hood and is evacuated. The exhaust is primarily CO2 produced from the anode consumption and hydrogen fluoride (HF) from the cryolite and flux. HF is a highly corrosive and toxic gas, even etching glass surfaces. The gases are either treated or vented into the atmosphere; the former involving neutralization of the HF to its sodium salt, sodium fluoride. The particulates are also captured and reused using electrostatic or bag filters. The remaining CO2 is usually vented into the atmosphere.
The very large electric current passing through the electrolytic cells generates a powerful magnetic field, and this can stir the molten aluminum with magneto-hydrodynamic forces in properly-designed cells. The stirring of the molten aluminum in each cell typically increases its performance, but the purity of the aluminum is reduced, since it gets mixed with small amounts of cryolite and aluminum fluoride. If the cells are designed for no stirring, they can be operated with static pools of molten aluminum so that the impurities either rise to the top of the metallic aluminum, or else sink to the bottom, leaving high-purity aluminum in the middle.
Aluminum smelters are usually sited where inexpensive hydroelectric power is available. For some European smelters, the electric power produced by hydroelectric plants in countries such as Norway, Switzerland, and Austria is transmitted by high-voltage power lines to such places as Denmark, Sweden, Germany, and Italy to be used by aluminum and magnesium factories. Since aluminum factories require nearly-uniform supplies of electric current, they make the most of nearly-constant supplies of electric power, and these are also available close to many hydroelectric power plants.
An alternate source of power, used by Icelandic smelters, is geothermal electricity, which Iceland has in abundance owing to its location on the Mid-Atlantic Ridge, but cannot use domestically owing to its low population. Iceland thus imports raw aluminum ore and uses the Hall–Héroult process as a means of exporting its electricity and thus efficiently exploiting the abundance of geothermal energy.
Impact
Aluminum is the most abundant metallic element on Earth but it is rarely found in its elemental state. It occurs in many minerals but its primary commercial source is bauxite, a mixture of hydrated aluminum oxides and compounds of other elements such as iron. It is converted to aluminum oxide, Al2O3 by the Bayer process and then to metallic aluminum by the Hall–Héroult process.
Prior to the Hall–Héroult process, elemental aluminum was made by heating ore along with elemental sodium or potassium in a vacuum. The method was complicated and consumed materials that were in themselves expensive at that time. This meant the cost to produce a small amount of aluminum in the early 19th century was very high.
Early aluminum was more costly than gold or platinum. Bars of aluminum were exhibited alongside the French crown jewels at the Exposition Universelle of 1855, and Emperor Napoleon III of France was said to have reserved his few sets of aluminum dinner plates and eating utensils for his most honored guests.
Production costs using older methods did come down, but when aluminum was selected as the material for the cap/lightning rod to sit atop the Washington Monument in Washington, D.C., it was still more expensive than silver.
New production based on the Hall–Héroult process, in combination with cheaper electric power, helped make aluminum (and incidentally magnesium) an inexpensive commodity.
This in turn helped make it possible for pioneers like Hugo Junkers to utilize aluminum and aluminum-magnesium alloys to make items like metal airplanes by the thousands, or Howard Lund to make aluminum fishing boats.
Castner process
The Castner process is a process for manufacturing sodium metal by electrolysis of molten sodium hydroxide at approximately 330°C. Below that temperature the melt would solidify, above that temperature, the metal would start to dissolve in the melt.
Fig. 6: Diagram of Castner process apparatus
Process
The diagram shows a ceramic crucible with a steel cylinder suspended within. Both cathode (C) and anode (A) are made of iron or nickel. The temperature is cooler at the bottom and hotter at the top so that the sodium hydroxide is solid in the neck (B) and liquid in the body of the vessel. Sodium metal forms at the cathode but is less dense than the fused sodium hydroxide electrolyte. Wire gauze (G) confines the sodium metal to accumulating at the top of the collection device (P). The cathode reaction is
2Na+ + 2e– → 2Na
The anode reaction is
2OH– → ½O2 + H2O + 2e–
Despite the elevated temperature some of the water produced remains dissolved in the electrolyte. This water diffuses throughout the electrolyte and results in the reverse reaction taking place on the electrolyzed sodium metal:
Na + H2O → ½H2 + Na+ + OH–
with the hydrogen gas also accumulating at (P). This, of course, reduces the efficiency of the process.
Van Arkel process
This method is employed to obtain ultra-pure metals. The impure metal is converted into a volatile compound while the impurities are not affected. The volatile compound is then decomposed electrically to get the pure metal. Other metals that can be purified this method are Zr, V, W, Hf etc.